No. 8814 



IC ^^14 



Bureau of Mines Information Circular/1980 




Valuation of Potash Occurrences 
Within the Nuclear Waste 
Isolation Pilot Plant Site 
in Southeastern New Mexico 



By Robert C. Weisner, Jim F. Lemons, Jr., 
and Luis V. Coppa 




UNITED STATES DEPARTMENT OF THE INTERIOR 



Information Circular 8814 

Valuation of Potash Occurrences 
Within the Nuclear Waste 
Isolation Pilot Plant Site 
in Southeastern New Mexico 



By Robert C. Weisner, Jim F. Lemons, Jr. 
and Luis V. Coppa 




UNITED STATES DEPARTMENT OF THE INTERIOR 
Cecil D. Andrus, Secretary 

BUREAU OF MINES 

Lindsay D. Norman, Acting Director 






Library 


of Congress Cataloging 


in Publication Data 






Weisner, Robert C 








V'alualioii of potash 


)a iirreiues within the nuclear waste 




isolation pilot plant site in southeastern New 


Mexico. 




(Iniormation circular— li 


ircau of Mines; 8814) 






BiblioKraphv: p. H(l 








Supi. of Doc. no.: I 28.27 








1. Potash deposits — New Mexico — (Carlsbad region. I. Lemons, Jim 
F., joint author. 11. Coppa, Luis V., joint author. IIL Title. IV. Series: 
United States. Bureau of Mines. Information circular; 8814. 




TN295.U4 [TN9191 622 


.08s [553'.636'0978942] 


79-607990 



For sale by the Superintendent of Documents, U.S. Government Printing Office 
Washington, D.C. 20402 



Stock No. 024-004-01959-0 



■^ 

^ 



PREFACE 

This report was prepared during the period 1976 through 1977 to serve a 
specific need of the Energy Research and Development Administration (ERDA), 
now a part of the Department of Energy. Subsequent broader interest in the 
report has demonstrated the need for publication. Some of the information 
presented on potash is now dated. The latest available statistics are published 
regularly by the Bureau of Mines in Mineral Industry Surveys, Mineral Com- 
modity Summaries, and Mineral Commodity Profiles. Readers needing current 
potash information can contact the Bureau of Mines commodity specialist: 

R. H. Singleton 
Bureau of Mines 
2401 E Street, NW. 
Washington, D.C. 20241 
Telephone (202) 634-1 190 






TERMS AND ABBREVIATIONS 

W/PP— Waste Isolation Pilot Plant. 

The WIPP site is the area within the boundaries of land proposed for with- 
drawal for the waste isolation facility. The site is subdivided into Zones I through 
IV. Zone I is the area of the proposed location of the isolation facility; Zones 
II through IV are located concentrically at approximately 1-mile intervals from 
center of the site. 

Ore zones is a term used to identify salt beds of minerals that locally contain 
potash minerals in commercial amounts. 

The Bureau of Mines study area includes lands in the WIPP site, and adjacent 
lands extending out for up to 2 miles (3 kilometers) from the site boundary. 

Mining Unit is a reasonable sized operating unit in terms of daily operating 
capacity and in lateral extent of workings for optimum ore handling, disre- 
garding any limits imposed by the current proposed WIPP site. Size limits also 
considered were (1) present and projected market size and (2) quantity of process 
water available. The unit values for the potash, determined for a Mining Unit, 
were then used to determine the total potash values within the WIPP site. 

K2O — by usage and for convenience, the concentrations of various com- 
mercial potassium minerals in ores, concentrates, and products are often stated 
as equivalent concentrations of potassium oxide (KjO). 



CONTENTS 



Page 

Preface _ i 

Terms and abbreviations ii 

Abstract 1 

Introduction _._ _ 1 

Acknowledgments 2 

Survey of the domestic potash industry _ 3 

History 3 

Economic background 4 

Market analysis and market projections 5 

Location and description of study area 9 

Geology of the WIPP area _____ 14 

Regional geology ____ 14 

Local geology _ 14 

Stratigraphy __._ _ 22 

Data collection and interpretation __. 23 

Ore zones and marker beds in the Salado Formation 23 

Potassium-bearing deposit estimation 23 

Evaluation of potash mineralization 25 

Current processing of potassium-bearing ores 25 

Sylvite 25 

Langbeinite __ 25 

Mixed ore _ 25 

Potassium sulfate _ 29 

Current impurities treatment 30 

Bench-scale metallurgical tests 30 

Sylvite flotation tests 31 

Sample AEC 7-5 32 

Sample AEC 8-10 ____ 33 

Langbeinite leach tests 33 

Sample AEC 8-4A 34 

Sample AEC 8-4B __ 34 

Sample AEC 8-4C 34 

Summary of metallurgical tests _ 35 

Mineralization in the WIPP site from U.S. Geological 

Survey data _ 35 

Financial analysis 51 

Methods used __ 51 

Cash flow estimates _._ _.__ 51 

Investment estimates 51 

Determination of commercial and subeconomic 

mineralization __ 52 

Taxes and royalties 52 

Federal corporation income tax 52 

New Mexico corporation income tax 52 

Resources, processors, and service tax 53 



Page 

Severance tax 53 

Property tax 53 

Rents and royalties on Federal leases 53 

Rents and royalties on State leases 54 

Estimation of current mine and mill capital and op- 
erating costs ___ 54 

Operations for WIPP study 54 

13,000-ton-per-day mine-mill 54 

7,000-ton-per-day mine-mill _._. 54 

8,500-ton-per-day mine-mill _ 58 

Mine capital costs ___ 59 

Mine operating costs 59 

Mill capital costs 63 

Mixed ore plant 63 

Sylvite ore flotation plant _ 63 

Sylvite ore crystallization plant 63 

Mill operating costs 63 

Utility (infrastructure) costs ._ 64 

Water 64 

Electric power _ 64 

Access to the area 65 

Estimates of hypothetical mining units in the WIPP 

site 66 

Mining unit A 66 

Mining unit A— 1 66 

Mining unit A-2 68 

Mining unit A-3 _ __ 69 

Mining unit B 69 

Mining unit B-1 ___ 69 

Mining units B-2 and B-3 70 

Mining unit C 70 

Mining unit C— 1 70 

Mining unit C-2 71 

Mining unit C-3 71 

Mining unit D 72 

Mining unit D-2 ____ 72 

Mining unit D-3 72 

Estimated mining unit capital investments and op- 
erating costs __ 72 

Summary and conclusions 75 

References.: 80 

Appendix A. — Assumptions and estimations used for 
the economic analysis of the hypothetical Mining 

Unit B-1 (MU B-1) _ 81 

Appendix B. — Financial evaluation of MU B-IA 84 



ILLUSTRATIONS 



1. Potash supply-demand relationships — 1974 __ ____ — - 7 

2. Annual trend for 20-year span of U.S. domestic potash prices __ ____ — 8 

3. Location map of the WIPP site _ _ __ — - -- -— 9 

4. Map of surface and mineral ownership, leases, amd lease applications in the study area _ 11 

5. Approximate location and size of WIPP zones within the study area _ ___ 12 

6. Major Permian geologic features in the region ____ ___ 15 

7. Stratigraphic column of consolidated rocks penetrated by site evaluation borings in Los Medanos area ___ 16 

8. Approximate location of cross sections at the WIPP site _.__ _ -- 17 

9. Approximate cross section through A-A' _ - 18 

10. Approximate cross section through B-B' _ - -- 19 

11. Approximate cross section through C-C _ — - - 20 

12. Approximate cross section through D-D' - - 21 

13. Stratigraphic column of the Salado Formation - -- Pocket 



Page 

14. Diagram of sylvite flotation section __ ___ ___ ___ 26 

15. Diagram of sylvite solution-crystallization section _ ___ ___ __ 27 

16. Diagram of mixed ore section __ ___ ___ ___ 28 

17. Diagram of a sulfate section ____ ___ ____ 29 

18. Composite map of mineralization in various ore zones at 8 and 14 percent KjO as langbeinite and sylvite, 

respectively _ ____ ____ _ __ 48 

19. Composite map of mineralization in various ore zones at 4 and 10 percent KjO as langbeinite and sylvite, 

respectively ___ ___ _ 49 

20. Composite map of mineralization in various ore zones at 3 and 8 percent KjO as langbeinite and sylvite, 

respectively __ ____ __ 50 

21. Generalized layout of 13,000-tpd surface plant ___ ____ ____ 55 

22. Generalized plan of room-and-pillar mining system showing typical ventilation flow _ _ 56 

23. Generalized plan of room-and-pillar mining system showing typical conveyor layout 57 

24. Diagram of conventional mining sequence ____ ____ __ 58 

25. Typical round dimensions for conventional mining _._ ____ 59 

26. Generalized layout of 7,000-tpd surface plant __ ___ ___ 60 

27. Generalized layout of 8,500-tpd surface plant ___ ___ 61 

28. Diagram of mining sequence using continuous miners __ ___ 62 

29. Approximate locations of mining units __ ___ ____ ___ 67 

30. Product availability and market prices at which potash tonnages become commercial assuming fixed production 

costs __ 76 

31. Approximate cost analysis of increased market prices at which potash tonnages become commercial __ 77 

A-1. System diagram of hypothetical Mining Unit B-1 operation, with approximate minerals flow through the plants 

and estimated gross revenues ____ __ _ 82 



TABLES 

Page 

1. Potash supply-demand relationships, 1965-76 _ 5 

2. World mine production and reserves _ 6 

3. Domestic potash producers, yearend 1976 _ 6 

4. Surface and mineral ownership 10 

5. Chemical analysis of core samples used in Bureau of Mines metallurgical test _ 31 

6. Approximate mineral content (wt-pct) of core samples used in Bureau tests 32 

7. Chemical assay mass balance of sylvite flotation: test on sample AEC 7-5 33 

8. Chemical assay mass balance of sylvite flotation: test on sample AEC 8-10 33 

9. Chemical assay mass balance of langbeinite leach: test on sample AEC 8-4A _ 34 

10. Chemical assay mass balance of langbeinite leach: test on sample AEC 8-4B .___ _ 34 

11. Chemical assay mass balance of langbeinite leach: test on sample AEC 8-4C 35 

12. Ore zone thickness and grade in test hole AEC-8 36 

13. Calculated mineral content of selected samples from potassium-bearing intervals with summation of percent 

KjO as ore mineral 37 

14. Summary of estimated mine capital investments _ _ 59 

15. Summary of estimated mine operating costs _._ 63 

16. Summary of estimated plant capital investments 63 

17. Summary of estimated annual mill operating costs ____ 64 

18. Comparison of fuel cost for a 3,000-tpd langbeinite wash plant 65 

19. Comparison of capital investment: coal versus natural gas in a 3,000-tpd langbeinite wash plant ___ 65 

20. Average K2O grade of sylvite and langbeinite deposits and total tonnage. Mining Unit A 68 

21. Mineralogical compositions of sylvite and langbeinite deposits. Mining Unit A _ _ _ 68 

22. Average KjO grade of langbeinite and total tonnage, Mining Unit B 70 

23. Mineralogical compositions of langbeinite deposits, Mining Unit B .___ 70 

24. Average KjO grade of sylvite and langbeinite deposits and total tonnage. Mining Unit C 70 

25. Mineralogical compositions of sylvite and langbeinite deposits. Mining Unit C 71 

26. Average K^O grade of langbeinite and total tonnage, Mining Unit D 72 

27. Mineralogical compositions of langbeinite deposits, Mining Unit D 72 

28. Summary of operating costs and capital investments for hypothetical Mining Units 73 

29. Potash mineral resource evaluation data in ERDA's WIPP site 73 

30. Mining Unit B-1 estimated potash values within the WIPP site that would be foregone 78 

31. Summary of amounts and values of potash mineralization in WIPP site _ 79 

32. Mining Unit product data and required market prices at which potash tonnages in the WIPP site become 

commercial at fixed production costs ___ _ 79 

33. Mining Unit price components per ton of weighted-average product price assuming fixed production costs and 

a 15 percent discounted cash flow rate of return — _ 79 

34. U.S. average market prices for potash products — f.o.b. plant ..._ 79 



VALUATION OF POTASH OCCURRENCES WITHIN THE 

NUCLEAR WASTE ISOLATION PILOT PLANT SITE IN 

SOUTHEASTERN NEW MEXICO 

by 
Robert C. Weisner,' Jim F. Lemons, Jr.,^ and Luis V. Coppa^ 



ABSTRACT 



Current production costs and market conditions in the potash industry of the 
Carlsbad area were studied to determine the potential values of the potash 
mineral resource that would be lost or foregone if the Waste Isolation Pilot Plant 
(WIPP) facility is constructed on the proposed site in that area. The purpose of 
the WIPP project is to investigate the possibility of developing a nuclear waste 
disposal plant in the salt formations at the site. Analyses were made of all potash 
deposits determined to be in the site. Mining and processing under the most 
favorable recovery systems were considered. Value determinations were based 
upon estimated operating and capital costs of current mine-mill operations in 
the Carlsbad area. This study was made for the Energy Research and Devel- 
opment Administration (ERDA) by members of the Federal Bureau of Mines 
Minerals Availability System staff. 

INTRODUCTION 

The use of nuclear fuels produces radioactive waste that must be stored or 
disposed of in an acceptable and safe manner. A safe storage location has been 
determined to be in thick salt beds. ERDA requested the U.S. Geological Survey 
to indicate the location of known salt beds so that a study could be made to 
determine those most suitable for disposal or storage sites. Geological Survey 
recommendations included the Los Medanos area of Eddy County, N. Mex. 

In September 1976 ERDA asked the Bureau of Mines to quantify and evaluate 
commercial potash minerahzation in a proposed WIPP site in southeastern New 
Mexico. This information was necessary as one element in an environmental 
impact assessment by ERDA of the proposed WIPP project investigating the 
possibility of developing a nuclear waste disposal plant in the salt formations at 
the site. 

The Bureau of Mines therefore conducted a study first to determine the 
amount and likely methods of recovery of existing potash deposits within the 



Mining engineer. MAS Division, Intermountain Field Operations Center, Bureau of Mines, Denver, Colo. 
Metallurgist, Minerals Availability Field Office, Bureau of Mines. Denver, Colo. 
Mining engineer. Minerals Availability Field Office, Bureau of Mines, Denver, Colo. 



WIPP site and immediate vicinity, and second to determine the value of com- 
mercial potash mineralization, present and future, within the site and study area. 
As part of the study, an analysis was made to determine if any of the potash 
mineral occurrences are commercially recoverable by existing mining and proc- 
essing techniques. This task was assigned to the Bureau's Minerals Availability 
System (MAS) personnel, since this group routinely conducts studies to deter- 
mine the availability and cost to the United States of minerals and metals. 

This report presents the details and results of the study prepared for ERDA. 
Included in the report are discussions of the geology and geography of the study 
area, its potash mining history, and current and projected market conditions in 
the potash industry. Prerequisite to the economic analysis, the Bureau of Mines 
obtained resource evaluation data from the Conservation Division of the U.S. 
Geological Survey (USGS) on measured and indicated categories of potash de- 
posits in the WIPP site; their grades and tonnages were then reevaluated and 
modified to ascertain grades and tonnages that could be recovered in mining 
and processing operations. 

The Bureau of Mines used criteria consistent with industry practice in pre- 
paring its economic feasibility studies; it employed a method of potash ore reserve 
calculations using engineering design and economic analytical procedures, in- 
cluding discounted cash flow, to determine the tonnage of minable potash ore 
that will yield an assumed, commercially acceptable (15 percent) rate of return 
on total capital investment. The mining and beneficiation systems evaluated were 
based on current extraction and processing technology in the Carlsbad district 
and were the least costly systems amenable to the ore to be mined. 

This analysis isolated and estimated the values that exist in the unmined potash 
mineralization in the WIPP site and therefore determined a cost chargeable to 
the WIPP facility as losses that would result from constructing the nuclear waste 
disposal plant. This cost is the sum of lost taxes, royalties, and bonus bid amounts 
that would otherwise be generated by development of the potash mineralization 
to the maximum commercial extent — potash values that would be foregone due 
to the closing of the site to future mining operations. 



ACKNOWLEDGMENTS 

The authors were assisted in this evaluation by the following personnel of the 
System Operations Group and the Domestic Evaluation Group, Minerals Avail- 
ability System, Intermountain Field Operations Center, Denver, Colorado: R. 
J. Minarik, W. L. Rice, R. L. Baer, L. B. Burgin, R. L. Davidoff, A. G. Hite, J. 
D. Lewis, C. M. Palencia, R. A. Salisbury, and R. C. Steckley. Data and assistance, 
especially helpful in preparing tables on worldwide and U.S. potash production, 
were given by R. H. Singleton, potassium commodity specialist of the Bureau 
of Mines Washington, D.C., headquarters office. Metallurgical testing was per- 
formed by personnel of the Bureau's Salt Lake City Research Center under the 
direction of Jerry L. Huiatt. 

The U.S. Geological Survey provided information on the amount, occurrence, 
distribution, mineralization, and grade of potash in the study area, and data 
interpretation, based on available information and data from ERDA exploration 
drill holes. Conversations with Charles L. Jones, who supervised the USGS core 
drilling within the WIPP site and authored the USGS report summarizing potash 
mineralization in the study area, were helpful in Bureau of Mines analyses. His 
geologic work, together with other USGS studies, provided the foundations for 
the geologic section of this report. 

Invaluable assistance was received from the potash industry in the Carlsbad 
area, utility companies, and State and local government offices. 



SURVEY OF THE DOMESTIC POTASH INDUSTRY 



HISTORY 

From early colonial times until 1860, the man- 
ufacture of potash from wood ashes constituted 
a significant chemical industry that met the 
needs of the United States and also provided an 
important commodity for export. This industry 
was seriously curtailed by the development of 
the Le Blanc process that provided a cheap 
source of sodium carbonate, which could be 
used instead of potassium carbonate for many 
industrial purposes. The end of this first potash 
industry came in 1861 when commercial mu- 
riate of potash was produced from deposits near 
Stassfurt in northern Germany. 

From 1861 to 1914, the potash industry was 
virtually nonexistent, and agriculture in the 
United States was almost totally dependent 
upon imported German potash (7).^ However, 
in January 1915, shortly after the onset of World 
War I, Germany placed an embargo on the ex- 
port of potash salts and the total U.S. supply 
was cut off. As a result, potash salts which pre- 
viously had sold at a normal price of $35.00 to 
$40.00 per ton ($38.58 to $44.09 per metric ton) 
were quoted at prices ranging from $350.00 to 
$425.00 per ton ($385.80 to $468.47 per metric 
ton) (6). These prices resulted in a burst of ac- 
tivity in the domestic potash industry. Methods 
of producing potassium compounds from kelp, 
wood ashes, lake brines, alunite, cement dust, 
sugar beet waste, blast furnace dust, and other 
sources were developed. By 1918 there were 128 
producers of potash compounds with a total an- 
nual production of approximately 55,000 tons 
(49,896 metric tons) of K20(i9). 

With the resumption of German imports fol- 
lowing the end of World War I, the domestic 
potash industry virtually collapsed again (20). 
By 1920 the American Trona Corp. (subse- 
quently American Potash and Chemical Corp. 
and now Kerr-McGee Chemical Corp.) was the 
only significant producer of potash salts in the 
United States. Between 1920 and 1930 the 
United States was again dependent upon for- 
eign imports for more than 80 percent of its 
potash requirements. 

"■ Underlined numbers in parentheses refer to items in the list of references 
preceding the appendixes. 



The Federal Government in 1924 authorized 
the Bureau of Mines and the U.S. Geological 
Survey "to determine location and extent of po- 
tash deposits in the United States" (5). By 1931, 
these agencies had identified, by core drilling, 
saline beds in the Permian Basin of Texas and 
New Mexico ranging in thickness from l-'/2 to 
more than 8 feet ( '/2 to more than 2 meters) and 
containing 9 to 14 percent K2O (18). 

Interest in the Permian Basin area had been 
generated by potash showings from oil explo- 
ration. The most significant of these showings 
was in the McNutt No. 1 oil test, drilled by the 
Snowden and McSweeney Co. in February 1925. 
Additional exploration by the Snowden and 
McSweeney Co. generated further interest and 
eventually resulted in the formation of the 
American Potash Co. This company continued 
the exploration efforts initiated by the Snowden 
and McSweeney Co. The minerals sylvite, po- 
lyhalite, langbeinite, and carnallite were found 
in many of the core tests, and the continuity of 
the bedded deposits became reasonably well es- 
tablished. In December 1929, the No. 1 shaft of 
the American Potash Co. was begun, and the 
first commercial production of potash from an 
underground mining operation began on March 
7, 1931. 

The name of American Potash Co. was changed 
to the U.S. Potash Co. in 1930. This company 
completed construction of its first potash refin- 
ery, a crystallization plant, and began produc- 
tion of muriate of potash on September 17, 
1932. Its second mine shaft was completed in 
June 1933, which increased mining capacity to 
more than 2,000 tons (1,814 metric tons) of ore 
per day. U.S. Potash Co. was merged and be- 
came U.S. Borax and Chemical Co. in 1956. The 
Carlsbad potash properties were than sold to 
U.S. Potash and Chemical Co., a subsidiary of 
Continental American Royalty Co., in 1968. 
Subsequent sales of this property were to Te- 
ledyne. Inc., in 1972 and to Mississippi Chemical 
Co. in 1974 (5). 

Potash Co. of America (now a division of Ideal 
Basic Industries, Inc.) began potash production 
in the Carlsbad area in 1934. This company used 
the same room-and-pillar mining methods as 
U.S. Potash Co., but its refinery consisted of a 



halite flotation process to separate the sylvite 
and halite minerals. This company has been in 
continuous production since 1934. Its refining 
process has been converted from halite flotation 
to the more economical potash flotation method 
now in almost universal use. 

Union Potash and Chemical Co. (now Inter- 
national Minerals and Chemical Corp.) began 
mining both sylvite and langbeinite in 1940 
(9). This refinery was the first commercial ap- 
plication of the potash flotation process (rather 
than halite flotation) for muriate of potash pro- 
duction. The langbeinite was refined by a leach- 
ing process. Commercial production of 
agricultural-grade potassium sulfate began 
shortly thereafter. Since the only known com- 
mercial deposits of langbeinite ore are in the 
Carlsbad area, this was the first commercial pro- 
duction of this mineral worldwide. 

During World War II, the production capa- 
bilities of the three Carlsbad potash companies, 
plus the production from American Potash and 
Chemical Co. at Trona, Calif., and Bonneville 
Potash, Ltd. (now Kaiser Aluminum and Chem- 
ical Corp.) at Wendover, Utah, prevented a se- 
rious shortage of potash in the United States. 
Potash production was classified as a defense 
industry, and appropriate priorities were pro- 
vided for personnel, materials, and equipment 
so that maximum production was maintained 
during the war years. 

Postwar expansion of the potash industry be- 
gan with the completion of the Duval Corp. (now 
a division of Pennzoil Corp.) plant near Carlsbad 
in 1951. This project was followed by the South- 
west Potash Co. (a wholly owned subsidiary of 
AM AX, Inc.) construction in 1952, and the Na- 
tional Potash Co. (a division of Freeport Min- 
erals Corp.) construction in 1957. All three of 
these operations used a potash flotation process 
to refine their sylvite ores. In 1964 Duval 
opened a new mine (Nash Draw) to produce 
langbeinite ore and increase its sylvite reserves. 
The Duval flotation plant was expanded at that 
time to include a langbeinite leaching process 
and to provide facilities for the production of 
potassium sulfate. 

Development of the first domestic under- 
ground potash operation outside the Carlsbad 
area was begun in 1961 in the Paradox Basin 
near Moab, Utah, by Texas Gulf Sulphur Co. 
(now Texasgulf, Inc.). Production of muriate of 
potash from this operation began in 1965, but 
mining problems made this operation uneco- 
nomic. In 1971 Texasgulf removed all equip- 
ment from the mine and started using another 
technique for potash recovery. Water is pumped 
into the mine from the nearby Colorado River 



and becomes saturated with potassium and so- 
dium chlorides; it is then pumped out of the 
"mine" to solar evaporation ponds. On evapo- 
ration, a mixture of sylvite and halite is depos- 
ited in the ponds. These salts are then harvested 
and refined in a conventional potash flotation 
plant to produce muriate of potash. 

The most recent domestic potash operation 
is Kermac Potash Co. (a division of Kerr-McGee 
Corp.), which began production of muriate of 
potash in 1965. Kermac is located a few miles 
east of the other six producers in the Carlsbad 
area. Because the characteristics of the Kermac 
ore made it very difficult to refine by flotation, 
a crystallization process was developed. This 
process makes a higher quality (K2O content) 
muriate product. 

During the 1960's a major expansion in potash 
production capacity occurred in Saskatchewan, 
Canada. As these large new operations came on- 
stream, the world supply of muriate of potash 
increased substantially and prices dropped. In- 
vestigation by the U.S. Government concluded 
that there was evidence of "dumping," and eco- 
nomic sanctions were proposed against some 
Canadian producers. Under the threat of these 
proposed economic sanctions, the Government 
of Saskatchewan in 1970 imposed production 
limitations and established minimum prices for 
Canadian potash. These actions by the Saskatch- 
ewan Government returned a degree of stability 
to the domestic potash industry (22). 

ECONOMIC BACKGROUND 

The world potash industry is regarded by 
some as an oligopoly (17). They contend there 
is a recognized interdependence in the market 
and high cross elasticities of demand. That is, 
from the buyer's viewpoint, potash is an undif- 
ferentiated product, equal in specifications from 
all suppliers, making market price virtually the 
only consideration in purchasing. Because po- 
tash is a comparatively high-bulk, low-value 
commodity, its marketing is affected signifi- 
candy by transportation costs that constitute a 
substantial part of the market price. Almost 50 
percent of U.S. consumption is in Illinois, Iowa, 
Indiana, Minnesota, Ohio, and Wisconsin. Be- 
cause of the transportation cost, Saskatchewan 
producers maintain an advantage over the New 
Mexico producers in these markets (1 7). 

Prior to World War II, price leadership was 
the coordinating mechanism for determining 
price and market share in the industry. Since 
World War II, structural changes in both the 
selling and buying markets have progressively 
Umited the feasibility of collusion among potash- 



producing firms. In fact, during the late 1960's, 
price cutting became widespread and the entire 
price structure collapsed, resulting in severe fi- 
nancial losses. The aggregate demand for po- 
tash is generally regarded as relatively inelastic, 
mainly due to the minor position of fertilizer in 
total farming costs. 

About 1960 "bulk blending" of fertilizers be- 
gan. Natural gas is used in petrochemical com- 
plexes as the key ingredient for producing 
ammonia that is combined directly with phos- 
phoric acid to manufacture nitrogen-phosphate 
compounds. These nitrogen-phosphate com- 
pounds are then blended with potash prior to 
being sold as fertilizers. The implementation of 
this process had a threefold impact on the po- 
tash industry. First, some petroleum companies 
in the fertilizer business, because of their natural 
gas product, integrated vertically and horizon- 
tally by acquiring financial interests in the po- 
tash industry. Second, long-term contractual 
agreements for potash purchases were negoti- 
ated; and third, the number of potential buyers 
of straight potash fertilizer was greatly reduced. 

Entry into the potash industry by a new pro- 
ducer is difficult because of the lack of exploit- 
able deposits, the requirement of technological 
expertise, large capital requirements, and a 
likely cost disadvantage to a new firm as com- 
pared to an established company. 

MARKET ANALYSIS AND MARKET 
PROJECTIONS 

During 1975, United States consumption of 
potash (KgO) decreased 17 percent to 5 million 
short tons (4.5 million metric tons)(table 1). This 



decrease, the first since 1961, was mainly due 
to general resistance by the agriculture industry 
to high fertilizer prices. Of total U.S. potash 
production, 95 percent is consumed by the fer- 
tilizer industry. 

This reversal proved short-lived, however, as 
1976 (estimated) consumption was up to 6 mil- 
lion short tons (5 million metric tons) repre- 
senting a 26-percent increase over the 1975 
figure and exceeding the previous (1974) high 
by more than 300,000 short tons (272,000 metric 
tons). Lower prices and supplier discounting 
during the summer largely accounted for this 
increase. 

In 1976, domestic production of potash was 
approximately 2.4 million short tons (about 2.2 
million metric tons), about 40 percent of do- 
mestic demand. By 2000, U.S. production is pro- 
jected to decline significantly, supplying less 
than 10 percent of the total U.S. consumption 
in that year (22). Domestic consumption is pro- 
jected to double to 12 million short tons (11 
million metric tons) of KgO. This increased pro- 
duction represents a growth rate of 2.9 percent 
per year (76), a decline from the previous 6.3- 
percent annual growth rate. The decline is based 
on the assumption of increasing physical limi- 
tations on potash consumption. Examples of 
these limitations include the decline of available 
land for farming, more efficient use of fertilizer, 
new technology in the use of fertilizer, and a 
possible decline in the foreign market for U.S. 
food products, particularly in the developing 
nations. Production and consumption rates pro- 
jected for 2000 could be changed by such de- 
velopments as higher foreign prices for potash, 
thereby making the deeply buried beds in Mich- 





TABLE L- 


— Potash supply-demand relationships, 

(Thousand short tons of KjO) 


1965-76 










1965 


1966 


1967 


1968 


1969 


1970 


1971 


1972 


1973 


1974 


1975 


1976"^^ 


World production: 
United States 


3,140 
12.060 


3,320 
12,739 


3,299 
14,054 


2,722 
15,145 


2,804 
16,394 


2,729 

17,284 


2,587 
19,358 


2,659 
19,401 


2,603 
21,695 


2,552 
23,516 


2,501 
24,922 


2,390 


Rest of world 


25,110 








Total 


15,200 


16,059 


17,353 


17.867 


19,198 


20.0 1 3 


21,945 


22,060 


24.298 


26,068 


27,423 


27,500 








Components of U.S. supply: 
Domestic mines 


3,140 

1,108 

295 


3,320 

1,491 

504 


3,299 
1,708 
690 


2,722 
2,166 


2,804 

2,332 

676 


2,729 

2,605 

392 


2,587 

2,766 

454 


2,659 

2,961 

428 


2,603 

3,587 

468 


2,552 

4,326 

206 


2,501 

3,736 

211 


2,390 




4,741 




619 








Total U.S. supply 


4,543 

504 

648 

3,391 


5,315 

690 

621 

4,004 


5,697 

863 

693 

4,141 


5,751 

676 

735 

4,340 


5,812 

392 

700 

4,720 


5,726 

454 

544 

4,728 


5,807 

428 

564 

4,815 


6,048 

468 

764 

4,816 


6,658 

206 

889 

5,563 


7,084 

211 

787 

6,086 


6,448 

619 

759 

5,061 


7,750 


Distribution of U.S. supply: 




Exports 


891 


Demand.. 




6,359 


U.S. demand pattern: 
Aericulture 


3,174 
217 


3,771 
233 


3,913 
228 


4,101 
239 


4,490 
230 


4,516 
212 


4,566 
249 


4,538 
278 


5,261 
302 


5,792 
294 


- 




Chemicals 






and 

nit (20 lb) of K2O, 
f.o.b. Carlsbad (av- 




Total U.S. primary dem 
Price: Cents per short ton u 
standard 60 percent muriate. 


3,391 
39 


4,004 
39 


4,141 

34 


4,340 
29 


4,720 
25 


4,728 
33 


4,815 
34 


4,816 
34 


5,563 
35 


6,086 
49 


5,061 
73 


6,359 
72 






'Estimate. 





























igan and North Dakota economic, or the failure 
of the domestic agriculture market to develop 
as anticipated. 

Estimated U.S. exports in 1976 increased 
about 15 percent from the 1975 figure to nearly 
900,000 short tons (8 1 0,000 metric tons) of KgO. 
North America produced about 25 percent of 
the world's potash in 1976. Estimated annual 
production capacity of potash in North America 
totals 10 million tons (9.1 million metric tons) 
of KgO, nearly three-quarters of which is in Sas- 
katchewan (table 2). 

About 82 percent of the 1976 domestic pro- 
duction was from the Carlsbad region in New 
Mexico; the balance was from Utah and Cali- 
fornia. Ten companies comprise the domestic 
industry. Nine companies have only one oper- 
ation, while one (Kerr-McGee) has a mine in 
New Mexico and a plant in California (table 3). 

About 15 percent of U.S. products are potas- 
sium sulfate or potassium magnesium sulfate. 
In 1975, U.S. production of these sulfates to- 
taled nearly 400,000 tons (360,000 metric tons). 
Sulfate compounds are produced both from 
langbeinite ore by two Carlsbad companies and 
from brines in Utah and California. However, 
the Carlsbad district is the only known source 
of commercial langbeinite mineralization in the 
United States. About one-third of the sulfate 
production is exported. 

Net potash imports have grown steadily to 
meet increasing U.S. consumption. In 1974, 96 
percent of the imports were from Canada, 1 
percent from Israel, the balance from the 
U.S.S.R., West Germany, and other countries 
(fig. I)(i5). 

Domestic potash reserves are estimated to be 
about 200 million short tons (about 181 million 
metric tons) of K2O recoverable at 1973 prices. 
These reserves include about 100 million short 
tons (about 90.7 million metric tons) in bedded 
deposits in New Mexico, 70 million short tons 
(64 million metric tons) in brines, and an esti- 
mated 30 million short tons (27 million metric 
tons) in Utah bedded deposits (14). 

In addition to the reserves, domestic potash 
resources include perhaps an additional billion 
or more tons of KjO, mostly in the extension of 
the Williston Basin southward into the Montana- 
North Dakota area and revised estimates of de- 
posits in the Paradox Basin in Utah. Large de- 
posits in Montana and North Dakota are now 
being studied by Kalium Chemicals, a subsidiary 
of PPG Industries, and others for possible re- 
covery by solution mining. Exploratory drilling 
was begun in 1976 in the Montana-North Da- 
kota area near the Canadian (Saskatchewan) 
border by two operators, Kalium Chemicals and 



TABLE 2. — World mine production and 
reserves 

(Thousand short tons of K^O) 





Production 




World mine production and reserves 


1975 


1976' 


Reserves 


United States: 
WIPP Site 



W 

W 



W 
W 


6 000 






Rest of United States 


1 00 000 






Total 


2,501 

5.992 

309 

2,298 

2,450 

789 

160 

506 

50 

12.368 


2.390 

5,400 

300 

2,300 

2,200 

700 

160 

500 

250 

13,300 


206 000 


Canada ___. 




Congo (Brazzaville) 


20.000 




100 000 






Israel and Jordan (Dead Sea) 


240 000 


Italy _. 




Spain ._ _ 




Other market economy countries 


250,000 






World total 


27,423 


27.500 








'Estimate. 

W Withheld to avoid disclosiiis company 


proprielarv 


data. 





TABLE 3. — Domestic potash producers, 
yearend 1976 

AMAX Chemical Corp. New Mexico. 

Duval Corp. (subsidiary of I'ennzoil) ,\ew Mexico. 

International Minerals & Chemical Corp. (IMC) .New Mexico. 

Great Salt Lake Minerals & (;hemical Corp. L'lali 

Kaiser Aluminum & Chemical Corp. Utah. 

Kcrr-McC.ee Corp. .New Mexico. 

Cililornia. 

Mississippi Chemical Corp. New Mexiio 

National Potash Co. (subsidiarN of Kreepori Miiier.iK 

Co.) New Mexico. 

Potash Company of America (subsidiar\ of Ideal Basic 

Industries) New Mexico 

Texasgulf liali. 



jointly by the Burlington Northern Railroad and 
C. F. Industries. It is reported both parties are 
planning large investments in solution mining, 
contingent upon the outcome of current feasi- 
bility studies. These deposits are about 7,000 to 
9,000 feet (about 2,134 to 2,743 meters) deep. 

During the first half of the 1970s (through 
1975) the price in constant dollars for North 
American potash approximately doubled. In the 
same period, constant-dollar prices for phos- 
phate and ammonia nearly quadrupled and tri- 
pled, respectively. The rate of these price rises 
escalated during this period and was particularly 
significant in 1974 and 1975. The average 1975 
quoted selling price for U.S. standard muriate 
grade potash was $77 per ton ($85 per metric 
ton) of K2O f.o.b. producers' plant site. The 
price increase trend for potash produced in the 
United States was reversed in June 1976 when 
the quoted price was lowered by the Potash Co. 
of America to $63 per ton ($69 per metric ton) 
of contained KgO f.o.b. Carlsbad, N. Mex. (fig. 
2). 

In 1962, the United States became a net im- 
porter of potash. This was due largely to the 
tremendous expansion in the Canadian potash 



World production 



East Germany 
3,158 



United States 
2,552 



France 
2,508 



U.S.S.R. 
6,700e 



Spain 
546 



Canada 
6,072 



4,164 



People's Republic 

of China 
375£ 



West Germany 
2,888 



Congo 

(Brazzavil le) 

350e 



Israel 
660e 



Italy 
230e 



Other 




Imports 
4,326 



Industry stocks 

1/1/74 

206 



Chile 
29e 



World total 
26,068e 







Industry s 


tocks 








12/3}/ H 
















U.S. 
7 


Supply 
084 




U.S. Demand 
6,086 




















Exports 










787 









Agriculture 
5,792 




SIC 287 


Chemicals 
294 





SIC 

Unit 
Note 



Key 

Estimate 

Standard Industrial Classification 
Thousand short tons of potash 
Includes only potassium muriate, 
potassium sodium nitrates, and 
potassium sulfate 



FIGURE 1. — Potash supply-demand relationships — 1974. 



industry. By 1968, the Canadians moved from 
a position of net importation and had captured 
45 percent of the U.S. market (24). On Decem- 
ber 19, 1969, the Treasury Department issued 
a finding that Canadian, West German, and 
French firms were selling muriate of potash 
(KCl) on the U.S. market at less than fair prices, 
in violation of the 1921 Antidumping Act. In 
1970, authority was given to the Tariff Com- 
mission (now called the International Trade 
Commission) to collect a dumping duty. How- 
ever, no duties have been collected because the 
Saskatchewan Government imposed restrictions 
on production, and prices of Canadian potash 
exported to the United States were increased. 
Nevertheless, firms seUing Canadian potash on 
the U.S. market have been subject to extensive 
reporting requirements, so that price differen- 
tials could be evaluated by U.S. Customs. Sub- 



sequent heavy demand and resulting price 
increases allowed the Saskatchewan Provincial 
Government to lift these controls in 1974. Most 
of the importers have been removed from the 
reporting list; several were dropped in 1974 and 
more on August 5, 1976. 

A complicated reserves tax was imposed by 
the Saskatchewan Government in October 1974, 
which, according to the Canadian Potash Asso- 
ciation, amounts to over 80 percent of gross 
profit after adding Federal, Provincial, and local 
taxes and levies. Members of the industry filed 
suit in Provincial court on the constitutionality 
of the reserves tax, and a lengthy litigation was 
promised by the Provincial Government through 
to the Canadian Supreme Court, if necessary. 

Legislation enabling acquisition of 50 percent 
or more of the Provincial potash industry was 
enacted by the Saskatchewan Government in 




FIGURE 2.— Annual trend for 20-year span of 
U.S. domestic potash prices. 



January 1976. The first unit, owned by Duval 
Corp., was acquired in October 1976 for $128.6 
million, and the second, owned by Sylvite of 
Canada, Ltd., was acquired in April 1977 for 
$144 million. These facilities represent about 18 
percent of the production capacity of the Sas- 
katchewan potash industry (14). 

The Antitrust Division of the U.S. Depart- 
ment of Justice met in early 1975 in Chicago to 
begin an investigation of the fertilizer industry. 
In June 1976, five U.S. producers were indicted 
and charged with restricting production and 
controlling prices in the United States and also 
with conspiring to coordinate United States and 
Canadian production for the control of prices 
of potash; they were also indicted on a charge 
of coordinating export of potash from the 
United States and the import of potash from 
outside North America into the United States. 
The potential fine is $50,000 per company, but 
more serious is the threat of customers bringing 
legal actions asking payment for damages sus- 
tained. Court trial began in Chicago in January 
1977; all five producers were acquitted in May 
1977. However, many class action civil suits seek- 
ing damages for overcharging are still pending. 



LOCATION AND DESCRIPTION OF STUDY AREA 



The proposed site considered by ERDA for 
a waste isolation pilot plant is 25 miles (40 kil- 
ometers) east of Carlsbad, N. Mex., and occupies 
about 29.6 square miles (76.7 square kilome- 
ters), with its center at the intersection of sec- 
tions 20, 21, 28, and 29, T 22 S, R 31 E (fig. 3). 
A study area of roughly 64 square miles (roughly 
166 square kilometers), including 34.4 square 



miles (89.1 square kilometers) outside the WIPP 
site, was examined to better understand the ge- 
ology and potash mineralization within the site. 
The study area consists of sections 1-2, 1 1-14, 
23-26, and 35-36 in T 22 S, R 30 E; all of T 22 
S, R 31 E; sections 1, 2, 11, and 12 in T 23 S, 
R 30 E; and sections 1-12 in T 23 S, R 31 E. 
Carlsbad is the nearest population center. The 





LEGEND 
Active mines 
Inactive shafts 
Refineries 
Abandoned refinery 



FIGURE 3.— Location map of the WIPP site. 



nearest potash operation is Duval Corporation's 
Nash Draw mine, about 2 miles (about 3 kilo- 
meters) west of the WIPP site boundary. Rec- 
ords indicate that four holes were drilled within 
the site for oil or gas; all were dry holes, and 
they have since been plugged and abandoned. 
Producing wells would be detrimental to ERDA 
objectives in that they could act as conductors 
of hydrocarbon gas or fluid and thus provide 
access of such hazardous materials to storage 
areas constructed in the site. Oil and gas pro- 
ducing formations are stratigraphically below 
the proposed site of the underground waste 
storage facility. The top of the salt beds is 750 
to 1,500 feet (229 to 457 meters) below the sur- 
face. The beds range from 1,500 to 2,000 feet 
(457 to 610 meters) thick and dip gently to the 
southeast. The area is stable tectonically with no 
known active faults. 

The Federal Government owns most of the 
land surface and mineral rights in the area, but 
there are some State and privately owned inter- 
ests. The ownership and areas of land involved 
are listed in table 4. 

Within the study area, all State lands and part 
of the Federal potash mineral lands are either 
leased or under lease applications. Figure 4 in- 
dicates the leaseholders' names and lease num- 
bers along with the names of the prospecting 
permit applicants and their application serial 
numbers. The size and location of the WIPP site 
with zones within the study area are depicted on 
figure 5. 

Access is afforded by State Highway 128 and 
from the north by U.S. Highway 180. If an all- 
weather road were constructed into the WIPP 
site, it probably would be built south from High- 
way 180. This highway is better suited to heavy 
haulage than Highway 128 and provides more 
convenient access to Carlsbad and other centers. 

A rail spur built into the area would probably 
be serviced by the Santa Fe Railroad as an ex- 
tension of its lines. A spur could be built as an 
extension of the Nash Draw mine or the Kermac 



TABLE 4. — Surface and mineral ownership 





In the study area 


In the WIPP site 




Acres 


Per- 
centage 


Acres 


Per- 
centage 


Federally owned surface and mineral 

rights. 
State owned surface and mineral 

rights. 
Privately owned surface and mineral 

rights. 
Privately owned surface and federally 

owned mineral rights. 
Privately owned surface and mineral 

rights except oil and gas federally 


35,440 

5,120 

80 

40 

280 


86.5 

12.5 

.2 

.1 
.7 


17,201.58 

1,759.49 

.00 

.00 

.00 


90.7 
9.3 


Total 


40,960 


100.0 


18.961.07 


100.0 



spurs (fig. 3). In either case, about 7 miles (about 
1 1 kilometers) of standard-gage rail line would 
be required. 

The water supply for a new potash-producing 
mine-mill complex is a critical item. Possible 
sources of water considered for potash refining 
are the Pecos River and aquifers in the Capitan 
Reef, Rusder, and Ogallala Formations. The 
best quality water found in the region is pro- 
duced from the Caprock area of the Ogallala 
Formation, 25 to 30 miles (40 to 48 kilometers) 
to the northeast of the study area. This source 
presently supplies the potash industry with 
about two-thirds of its water and is being used 
increasingly by the industry. The water quality 
is good, ranging from 500 to 600 parts per mil- 
lion total dissolved solids. 

This study assumes that Caprock area-Ogal- 
lala Formation water will be available for a new 
refinery. Adequate water rights are available 
and well site leases are obtainable on most of the 
Caprock field. The producing area is owned 
largely by the State of New Mexico. Present wells 
are 200 to 250 feet (61 to 76 meters) deep and 
are spaced on about '/4-mile (0.4-kilometer) cen- 
ters. Each well, if properly designed and devel- 
oped, can produce 200 to 250 gallons per 
minute (756 to 945 liters per minute) for 30 to 
40 years. It is estimated that a pipeline from the 
Caprock field to a new refinery in the study area 
would be 30 miles (48 kilometers) long. 

Natural gas for a potash refinery could be 
supplied by the Gas Co. of New Mexico. This 
company presendy supplies the Kermac refin- 
ery using a 6-inch (15-centimeter) line, which 
has sufficient capacity to supply an additional 
refinery and could be extended into the study 
area. 

Electric power is supplied by the Southwest- 
ern Public Service Co. A typical load for a potash 
mine and refinery is estimated to be about 6,000 
kilowatts, with usage at approximately 4 million 
kilowatt-hours per month. All power supplied 
to the potash industry comes over 69-kilovolt 
powerlines. An exisdng 69-kilovolt line to the 
Kermac operation, capable of handling the 
added capacity required for a new mine and mill 
complex, could be extended into the study area. 

Loving and Carlsbad are the towns nearest to 
the study area, located about 20 and 28 miles 
(32 and 45 kilometers) respectively, from its cen- 
ter. Total population of the two towns is about 
28,000, with about 20,000 people living in Carls- 
bad. 

Unskilled and semiskilled labor generally is 
available in the area. Trained miners and other 
skilled workers are presendy employed in the 



10 



R30E 



R3IE 




I I Potash leases, lease holders and lease 
numbers 

I I Potash prospecting permit application, 
applicants name and application number 

I I Federal surtace and mineral rights 

\''''\ State surface and mineral rights 

^ ' J Private surtace some with mineral rights 

^— Proposed WIPP site outline 

Zone boundary 

Lease boundary 



FIGURE 4. — Map of surface and mineral ownership, leases, and lease applications in the study 

area. 



11 



R30E 



R3IE 




I I Federal surface and mineral rights 

\.]]'/^ State surface and mineral rights 

['.J Private surface and mineral rights 

F^ Private surface, all mineral rights owned 
^^^ by Federal Government 

p^ Private surface and mineral rights, except 
^'^^ oil and gas federally owned 

—^Proposed WIPP site outline 

Zone boundaries and areas provided by ERDA 



I 


- 58 acres 


n 


- 1,889 acres 


m 


- 6,201 acres 


m 


- 10,812 acres 



FIGURE 5. — Approximate location and size of WIPP zones within the study area. 



12 



potash industry; therefore, a new operation all construction sites and pond sites must have 

would require training programs for new em- archeological surveys. Once surveys have been 

ployees. conducted, salvage operations of the important 

Present Federal procedures require that ar- sites must be conducted by a certified archeol- 

cheological surveys be made along all pipelines, ogist. These archeological su: veys and excava- 

powerlines, access roads, and railroad rights-of- tions will add to our knowledge of the cultural 

way granted by the Federal Government; also, heritage of the area. 



13 



GEOLOGY OF THE WIPP AREA 



REGIONAL GEOLOGY 

The WIPP site is near Carlsbad, N. Mex., in 
the western half of two adjoining structural bas- 
ins of Permian age. The two basins, the Dela- 
ware Basin on the west and the Midland Basin 
to the east, together comprise the much larger 
regional Permian Basin (21). The Delaware 
Basin occupies an area in southeastern New 
Mexico and west Texas roughly 135 miles 
(roughly 217 kilometers) long by 75 miles (121 
kilometers) wide (fig. 6). The basin is nearly sur- 
rounded by the large horseshoe-shaped Capitan 
Limestone that opens to the south. This reef 
grew on shallow platforms to the east and west 
and on the shelf area north of the Delaware 
Basin. Growth of the reef probably contributed 
to formation of the embayment in which potash 
salts were deposited. Capitan Limestone extends 
in subsurface eastward from Carlsbad to Hobbs 
and thence southeastward along the Central 
Basin Platform into Texas. To the south and 
west of Carlsbad, the reef is exposed as El Cap- 
itan Peak and forms a portion of the Guadalupe 
Mountains. 

Before Permian time the basin area was sub- 
merged, and a thick section of sedimentary 
rocks was deposited. Toward the end of Per- 
mian time, the reef growth was halted due to an 
influx of high-salinity seawater. The evaporite- 
bearing Castile, Salado, and Rustler Formations 
were then deposited on top of several thousand 
feet of Lower Paleozoic sedimentary rocks (fig. 
7). The evaporite sequence was followed by the 
deposition of terrestrial sedimentary rocks known 
as the Dewey Lake Redbeds. Later terrestrial 
sand, clay, and sandstone were deposited, and 
in part eroded, into Quaternary time. Today, 
much of the land surface is covered by caliche 
and low-lying sane dunes (1). 

Several economic minerals are found in the 
area. Significant gas and oil are produced from 
the Pennsylvanian age Strawn, Atoka, and Mor- 
row Formations. Less important hydrocarbon 
production is derived from scattered reservoirs 
in the Middle Paleozoic section in the Delaware 
Basin. Some sulfur has been produced from an- 
hydrite caprock material in the Rustler and Sa- 
lado Formations in Culbertson County, Tex., 



and potash minerals are mined from several 
mineral-bearing zones in the Salado Formation. 

LOCAL GEOLOGY 

Geologic information on the WIPP area has 
been compiled largely by Charles Jones {10-12) 
and other personnel of the U.S. Geological Sur- 
vey (2, 4, 8). Lower and Middle Paleozoic rocks 
do not play an important role in the WIPP site 
geology and are not pertinent to this report; 
only the Permian and younger rocks are briefly 
discussed for this economic analysis. 

Thick sections of salt at the top of the Salado 
Formation and anhydrite within the Rustler 
Formation have been dissolved and removed by 
ground water action, resulting in the creation 
of many sinkholes and a general lowering of the 
land surface over much of southeastern New 
Mexico. The most pronounced topographic fea- 
ture near the WIPP area is Nash Draw, a depres- 
sion that contains several sink areas within its 
boundaries (25). The most pronounced depres- 
sion within Nash Draw is a salt lake called La- 
guna Grande de la Sal. The Salado Formation 
does not crop out in the area. 

The Rustler Formation crops out in several 
places west and northwest of, but not within, the 
WIPP area. The Dewey Lake Redbeds are ex- 
posed in Nash Draw and along the western pe- 
rimeter of the area. Younger rocks in the area 
include alluvial bolson deposits and windblown 
sand (dunes). 

Structural deformation in the Permian rocks 
is limited to a gently eastward-dipping mono- 
cline with some minor flexures as illustrated in 
cross sections (figs. 8-12). A few collapse struc- 
tures can be found in the evaporite sections. 
Many pre-Permian structures are reflected in 
the overlying beds. Such a feature is a localized 
structural trough opening southeastward from 
the southeast corner of Sec 8, T 22 S, R 31 E. 
This trough shows up in the 1 1th and lower ore 
zones and extends for more than 4 miles (6 kil- 
ometers). The cause of such a structure is un- 
known, but it is hypothesized that the trough 
could be the expression of an ancient drainage- 
way or collapse feature. The trough appears in 
east-west cross section B-B' through the north- 



14 




FIGURE 6.— Major Permian geologic features in the region. Source: 



(22). 



15 



Santa 

Rosa 

Sandstone 



Dewey 

Lake 

Redbeds 



APPROXIMATE 

THICKNESS 

FEET (METRES) 



330 (100) 



(113) 



1,100 (335) 



>1,000 (>305) 



GENERAL CHARACTER 



Sandstone Interbedded 
stone 



Slltstone and very fine grained 
sandstone 



Anhydrite (gypsum) interbedded 
with dolomite, siltstone, and 
sandstone 



Rock salt interbedded with 
anhydrite, glauberite, silty 
sandstone, and a variety of 
potassium-bearing rocks 



Anhydrite and seams of rock salt 



EXPLANATION 



ESI 

Anhydrite and (or) 
other sulfate rock 



§1 
v c • 



s*s3 s: 

•H U -H 1-1 M 

m3 n > 3 

O H -H 0) W 
•H to tH kl 

« rH 

01 Z kl li CO 

U U Q. O U • 
Bm -H « 

• O a -a tB u 

« -H « o 3 

s a-H o « 

0(0 

a c X 01 

2$4i g 

H XI > B 



FIGURE 7. — Stratigraphic column of consolidated rocks penetrated by site evaluation borings sunk 

in Los Medanos area. 



16 




LEGEN D 



' — ^ '-^— ' Natural gas pipeline 
A ^A' Approximate lines of cross sections 



© Potash drill holes 

B ERDA potosh drillholes 

I I Federal surface and mineral rights ^—— Proposed WIPP site outline 

\^^[['\ State surface and mineral rights Zone boundary 

d Private surface some with mineral rights 



FIGURE 8. — Approximate location of cross sections at the WIPP site. 



17 



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21 



ern area. Structural features related to adorning 
effect from the possible hydration of gypsum 
are found several places outside the area. There 
are no known active faults within the WIPP site 
or study area. Rocks penetrated by drilling in 
the study area are from Late Permian to Late 
Triassic in age. 

STRATIGRAPHY 

The Upper Permian System in the WIPP area 
includes the Ochoan Series, which consists of 
four rock salt, anhydrite, and siltstone units; the 
Castile, Salado, and Rustler Formations; and the 
Dewey Lake Redbeds. The Salado Formation 
contains all the potash mineralization of eco- 
nomic importance (fig. 13, in pocket). The 
Ochoan Series overlies shaly siltstone and as- 
sociated limestone of the Guadalupian Series 
(Lower and Upper Permian) and is unconform- 
ably overlain by rocks ranging in age from Late 
Triassic to Middle Pleistocene. 

The Castile Formation is the deepest and old- 
est stratigraphic unit penetrated by test hole 
drilling in the WIPP site. In the ERDA No. 9 
borehole at the center of the area (fig. 8), the 
top of the formation was penetrated at a depth 
of 2,836 feet (864.4 meters). Thickness of the 
formation is estimated to be between 1,280 and 
1,740 feet (390 and 530 meters). Anhydrite is 
the main constituent along with a few interbeds 
of rock salt and a small amount of calcitic lime- 
stone. The upper contact of the Castile For- 
mation is conformable with the overlying Salado 
Formation where there is a lateral and vertical 
gradation from anhydrite to rock salt. 

The top of the Salado Formation in test hole 
ERDA No. 9 is at a depth of 860 feet (262 me- 
ters). No solution residue is at the top of the 
bed, indicating that it has not been affected by 
ground water as it has in other areas to the west. 
The formation thickness at this location is 1,976 
feet (602.3 meters). 

The Salado Formation is divided into three 
members: lower; middle or McNutt potash 
zone; and upper. The members are similar in 
lithology but differ in potash mineral content. 
As it exists at test hole ERDA No. 9, the top of 
the lower member is at a depth of 1,741 feet 
(530.7 meters), and the member is 1,095 feet 
(333.8 meters) thick. Information available shows 
that this member consists mainly of rock salt with 
minor amounts of anhydrite, polyhalite, and 
glauberite. 

The middle member, the McNutt potash 
zone, contains the currently economic potash 
minerals, langbeinite and sylvite. Mineralization 
in commercial concentrations is not present in 
test hole ERDA No. 9. The top of the McNutt 



potash zone is at a depth of 1,362 feet (415.1 
meters), and the zone is 379 feet (1 15.5 meters) 
thick. It is comprised of rock salt with interbed- 
ded polyhalite, minor anhydrite, and clay. The 
normal potash-bearing zone, barren at this 
point, is identified by a marker bed of anhydrite 
near the base and by a thin seam of silty sand- 
stone near the top. 

The upper member is 520 feet (158 meters) 
thick; the top is at a depth of 860 feet (262 me- 
ters). This member is mainly rock salt with a few 
interbeds of polyhalite, anhydrite, and brown 
sandstone. The upper boundary of the Salado 
Formation is a sharp but conformable contact 
between rock salt and siltstone of the overlying 
Rustler Formation. 

The top of the Rusder Formation is 550 feet 
(168 meters) below the surface, and the for- 
mation is 310 feet (94.5 meters) thick. The unit 
is composed of interbedded anhydrite, dolo- 
mite, salt, and fine-grained sandstone (fig. 13). 
The permeability of four beds in the Rustler 
Formation indicates that they may function as 
aquifers. These zones are a 24-foot (7.3-meter) 
dolomite section located at a depth of 608 feet 
(185 meters), a 15-foot (4.6-meter) anhydrite 
bed at a depth of 680 feet (207 meters), a 25- 
foot (7.6-meter) section of Culebra Dolomite at 
a depth of 714 feet (218 meters), and a 12-foot 
(3.7-meter) thickness of clay and minor silt 758 
feet (231 meters) below the surface. 

A distinct reddish-brown mudstone marks the 
sharp, unconformable contact of the Rustler 
Formation with the overlying Dewey Lake 
Redbeds. The Dewey Lake Redbeds comprise 
a red-colored silty unit located 63 feet (19 me- 
ters) below the surface. It is 487 feet (148 meters) 
thick and is present throughout the WIPP site. 
The Dewey Lake Redbeds have been eroded in 
post-Permian time and vary greatly in thickness; 
in the WIPP site these beds occur as a repeated 
sequence of siltstone and very fine-grained 
sandstone. 

The Triassic Santa Rosa Formation, consisting 
of interbedded sandstone and mudstone, un- 
conformably overlies the Dewey Lake Redbeds. 
Nine feet (3 meters) of medium-grained, friable 
sandstone is present. The top of the formation 
is at a depth of 54 feet (16 meters). 

Overlying the Santa Rosa Formation is the 
Gatuna Formation, which consists of 27 feet (8.2 
meters) of silty, calcitic sandstone. 

The Gatuna Formation is overlain by 5 feet 
(1.5 meters) of Pleistocene Mescalero Formation 
caliche, which is covered by dune sand. The lat- 
eral extent of the Gatuna and Mescalero For- 
mations is uncertain in the WIPP area because 
of erosion and mantling by dune sands. 



22 



DATA COLLECTION AND INTERPRETATION 



The U.S. Geological Survey recommended, as 
one of several sites suitable for a waste storage 
location, the use of thick salt beds in the lower 
member of the Salado Formation (fig. 13). Ex- 
ploratory holes were drilled to confirm the lith- 
ology, structure, mineralogy of the salt-bed 
sequence, and grades and amounts of potash 
minerals in the site area. For this purpose, 21 
holes were drilled on about 1-mile (about 1.6- 
kilometer) centers, and the cores were analyzed. 
Conventional rotary drilling methods were used 
to penetrate to the top of the McNutt potash 
zone. When this horizon was reached, the drill- 
ing fluid was changed to saturated brine to pre- 
vent dissolution of the minerals and to allow for 
good core recovery. The interval was cored 
through the McNutt potash zone into the under- 
lying lower member of the Salado Formation. 
The cores were visually logged by a geologist 
and splits were sent to laboratories for chemical 
analyses. The analytical data were used for pre- 
liminary estimates of tonnage and grade of po- 
tash occurring in the WIPP site. 



devised a system for identifying the strati- 
graphic units in the Carlsbad district that seg- 
regated the potash-bearing rock units into 
economic and noneconomic types. The calcium 
sulfate rocks carried no economic values, whereas 
the mineralized chloride or rock salt beds did. 
A system of marker beds was devised to number 
and identify the noneconomic calcium sulfate 
beds within the Salado Formation, beginning 
with marker bed 100 near the top and ending 
with marker bed 143 near the base. 

The salt beds of economic importance were 
found to occur in the middle member of the 
Salado Formation now known as the McNutt 
potash zone. Eleven mineralized salt zones were 
identified which were numbered from ore zone 
number 1 near the base of the member to ore 
zone number II near the top (fig. 13). A 1960 
open-file report by Jones, Bowles, and Bell pro- 
posed these new numbering systems for potash 
well logging and stratigraphic correlation pur- 
poses (13). The method was accepted by indus- 
try and is currently being used. 



ORE ZONES AND MARKER BEDS IN THE 
SALADO FORMATION 

During the late 1920's, the U.S. Geological 
Survey and the Bureau of Mines drilled the 
Carlsbad potash area. At that time, there was no 
geologic type section or nomenclature to delin- 
eate stratigraphic units or mineralized zones. 

As the drill samples were analyzed, chemists 
recognized various stratigraphic intervals con- 
taining high K,jO (potash) values. The potash 
usually occurred in argillaceous salt beds sepa- 
rated by calcium sulfate beds composed of an- 
hydrite and/or polyhalite and other barren rock 
salt beds. About 40 individual beds within the 
Salado Formation that contained some percent- 
age of K2O were numbered from the top of the 
formation downward in an effort to trace stra- 
tigraphic units and mineralized zones. Each po- 
tash company working in the area used a 
different type of nomenclature varying from let- 
ters to numbers or combinations of both to iden- 
tify beds and mineralized zones. 

In the late 1950's, the U.S. Geological Survey 



POTASSIUM-BEARING DEPOSIT 
ESTIMATION 

A triangular method to calculate potash ton- 
ages, using standards of a minimum bed thick- 
ness of 4 feet (1.2 meters) of 4 percent K2O as 
langbeinite [4 feet (1.2 meters) x 4 percent = 
16 feet-percent) (4.8-meter-percent)], or 4 feet 
(1.2 meters) of 10 percent K2O as sylvite [4 feet 
(1.2 meters) x 4 percent = 40 feet-percent (12 
meter- percent)], or the equivalent minimum 
feet-percentage product if thickness was less 
than 4 feet (1.2 meters). In other words, the 
thickness could be reduced below 4 feet (1.2 
meters) if the grade were high enough to meet 
the above minimum feet-percentage criteria. 
Triangles drawn to connect adjacent test holes 
peripheral to the drilling network were pro- 
jected outward for distances considered prudent 
based on geologic interpretation. The rule of 
linear distribution was used to determine cutoff 
points between holes having insufficient thick- 
ness and/or grade and holes having mineral con- 
centrations at or above cutoff grade. In a few 



23 



instances, a circle of influence with a '/.'-mile 
(0.8-kilometer) radius was established around an 
outlying hole of significant grade when no other 
holes existed within a distance of l-V-i to 2 miles 
(2.4 to 3.2 kilometers). 

The areas thus established were weight-av- 
eraged by mineralized bed thickness and grade 
to determine an average grade-thickness value. 
The areas were then measured by polar plani- 
meter, and the results multiplied by the average 
bed thickness to determine volumes that were 
converted to short tons. The average thickness 



and grade within each triangle were multiplied 
by the area of the triangle, and these numbers 
were divided by the sum of the areas to achieve 
a weight-averaged grade and thickness. These 
data were then used to determine recoverable 
tonnage based on data supplied by the U.S. Geo- 
logical Survey. The weight-averaged grade and 
thickness, tonnages, values, or zone elevations, 
and potash-bearing areas were plotted on base 
maps. This method of ore reserve estimation is 
suitable for uniformly layered sedimentary de- 
posits. 



24 



EVALUATION OF POTASH MINERALIZATION 



The following is a discussion of current flo- 
tation and leaching technology, the expected 
effect of impurities, and a discussion of the tests 
made by the Bureau of Mines. 

CURRENT PROCESSING OF POTASSIUM- 
BEARING ORES 

Currently, ores containing sylvite (KCl), lang- 
beinite (K2S04-2MgS04), and mixed-sylvite-and- 
langbeinite minerals are being mined and proc- 
essed in the Carlsbad area. The ores are up- 
graded by heavy media, flotation, leaching, and 
crystallization techniques that separate the de- 
sired potash minerals from halite, clays, slimes, 
and other mineral impurities. 

Sylvite 

Sylvite ores can be beneficiated by either flo- 
tation or solution-crystallization techniques. Both 
methods are being employed in the Carlsbad 
area. In a flotation process, sylvite ores contain- 
ing between 13 and 23 percent K^,0 equivalent 
are crushed, deslimed, and floated in a series of 
pneumatic flotation cells after conditioning with 
selective collectors and depressive reagents. Fig- 
ure 14 is a generalized schematic diagram of this 
process. 

The conditioners often include an amine col- 
lector, which makes the potassium chloride hy- 
drophobic; a blinder which depresses slime 
flotation; and an alcohol which acts as a frothing 
agent. The floated sylvite mineral is then dried, 
sized, and stored for market consumption under 
the product name muriate of potash (60 percent 
KgO). Fines separated in the sizing operation are 
either compacted and sold or are used as feed 
to a potassium sulfate operation that will be de- 
scribed later. Brine and tailings from the flo- 
tation operation are separated, and tailings are 
discarded. The brine may then be directly re- 
turned to the wash desliming circuit, or it may 
first go to a crystallization circuit where more of 
the dissolved sylvite is removed. Typical KCl 
recoveries of 80 to 87 percent are achieved with 
such a sylvite flotation process. 

When the clay and slimes impurities in sylvite 
ores increase to the range of more than 3.5 to 



4 percent, extensive mechanical desliming is re- 
quired, or the sylvite flotation recovery de- 
creases significantly. To avoid these difficulties 
and other liberation problems, one company is 
currently processing high-clay sylvite ore bv so- 
lution and crystallization methods. A schematic 
diagram of this process is shown in figure 15. 
The sylvite in the crushed and mechanicallv des- 
limed ore is leached in hot [185° to 200° F (85° 
to 93.3° C)] unsaturated brine. The svlvite 
product, muriate, is crystallized from the preg- 
nant leach liquor by vacuum cooling, then dried 
and stored for market. Typical KCl recoveries 
of 80 to 85 percent are achieved. 

Langbeinite 

Figures 16 and 17 illustrate a generalized be- 
neficiation method for langbeinite ores. These 
ores, typically containing 7.5 to 10 percent K,,0, 
are crushed and sized. The impurities are then 
water-leached from the less soluble langbeinite. 
Then, the lapgbeinite is dried, sized, and stored 
for market consumption under some form of 
the name sulfate of potash-magnesia (22 percent 
KgO). Typical langbeinite mill recoveries are 85 
to 90 percent. Fines from the sizing operation 
are used as feed to a potassium sulfate opera- 
tion. 

Mixed Ore 

Langbeinite may occur associated with sylvite 
in amounts such that both minerals can be re- 
covered. Currently, a mixed ore of sylvite (8 to 
10 percent KjO) and langbeinite (2 to 3 percent 
KjO) is being beneficiated in the Carlsbad area. 
A schematic diagram of this operation is shown 
in figure 16. After crushing, sizing, and me- 
chanical desHming of the ore, the two potassium 
minerals in the coarse (plus 20 mesh) ore are 
separated by a two-stage, magnetite-heavy-me- 
dia separation process. 

The denser langbeinite is separated as the sink 
product from the lighter sylvite and halite min- 
erals in the first heavy-media stage. In the sec- 
ond stage, the media specific gravity is readjusted 
and sylvite is separated as the float product from 
the halite. The separated sylvite joins the mill 



25 





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Brine 
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Section 8 


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26 



Ore 

X 



Crush 
Section 



Brine 



-3 mesh 



^ Deslime 
Section 2 



Brine 



Solids 



Hot Leach 
Section 3 



Tail ings 



Tailings 



Brine 



Debrine 
Section 4 



Leach 

Clarification 

Section 5 



Solution 



Tailings 
discard 



Tailings 
discard 




Brine 



Crystallization 
Section 7 



Crystal lization 
Debrining 
Section 8 



Muriate 

Dry, Size, 

Compact, Storage 

Section 9 



Key 

Section identifications 
refer to equipment lists 
for specific tonnage 
throughput 
FIGURE 15. — Diagram of sylvite solution-crystallization section. 



27 



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1 £ 




Langbeinite 
dry, size, 

storage 
Section 16 






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Langbeinite 
Flotation 

Section 17 


fo-^ 




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Section 15 


1 


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Coarse 

Flotation 

Section 14 


















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1 


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Magnetite-Heavy 

media input 

recovery and 

specific gravity 

adjustment circuit 






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heavy media 
separation 
Section 14 




Muriate 
sizing 

Section 8 


i 


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sizing 

Section 9 






















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Middlings 
Rescreening 

Section 6 




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Section 1 
1 


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Section 2 


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Section 3 




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Rougher 
Flotation 
Section 4 


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Dry, Size, 








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Cleaner 
Flotation 
Section 5 


ro ro c • I 


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otation tailings 

Section 12 


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28 



feed ore fines (minus 20 mesh) in a standard 
sylvite flotation circuit. The floated sylvite is 
dried, sized, compacted when necessary, and 
stored for market consumption as muriate of 
potash. The flotation tails are separated from 
the brine and combined with the langbeinite 
fines recovered from the heavy-media opera- 
tion. These langbeinite fines are then floated, 
dried, and stored for use as feed to a sulfate 
operation. The langbeinite sands recovered 



from the heavy-media operation are dried, 
sized, and stored for market. Typical recoveries 
of 60 to 70 percent langbeinite and 70 to 80 
percent sylvite are achieved. 

Potassium Sulfate 

Langbeinite product fines are typically com- 
bined with the necessary amount of sylvite prod- 
uct fines to produce the marketable product 





Langbeinite Fines 






Muriate Fines 










\f 




Pulverizing 
Section 1 




Dissolution 

and 

Clarification 

Section 2 




















r 










Sulfate 
reaction 

Section 3 






















Mixed 
salts 




\ 


f 






Debrine 
Section 4 


Brine _ 


Evaporation, 

crystallization 

and filtration 

Section 5 














1 




Sulfate 
dry, size, 

storage 
Section 6 




Waste liquor 
discard 



Section identifications refer to 
attached equipment lists for 
specific tonnage throughput 

Langbeinite and muriate fines are concentrate 
products from the langbeinite leach and sylvite 
flotation circuits. 

FIGURE 17. — Diagram of a sulfate section. 



29 



potassium sulfate; figure 17 is a schematic dia- 
gram of this operation. In this base exchange 
reaction, dissolved muriate reacts with pulver- 
ized langbeinite to form soluble potassium sul- 
fate and magnesium chloride. The potassium 
sulfate is then crystallized from the solution 
under controlled conditions as the temperature 
decreases. The potassium sulfate product (50 
percent K^O) is then separated from reaction 
liquor, dried, and stored for market. Mixed po- 
tassium salts remaining in the reaction liquor, 
including potassium chloride, are then re- 
covered by vacuum crystallization and are re- 
turned to the feed input. Then the brine, which 
still contains the magnesium chloride and other 
noncrystallized salts, is discarded. Typical re- 
coveries of potassium sulfate range between 70 
and 80 percent. 

Current Impurities Treatment 

Other potassium-containing minerals and clays 
associated with the sylvite and langbeinite ores 
can drastically affect recovery and concentrate 
grades. The impurities, if not removed, can 
make the product not marketable, and their re- 
moval can greatly increase the capital invest- 
ments needed to recover a marketable product. 

The effect of clays and slimes in sylvite re- 
covery has already been discussed as one of the 
major processing difficulties. Companies proc- 
essing high-clay ores by flotation have found it 
necessary to increase desliming and brine clar- 
ification circuits and to increase the addition of 
flocculant and blinder reagents. Even with these 
modifications, as much as 7 to 8 percent greater 
loss of K2O has been noted in processing some 
high-clay ore. New processing techniques aimed 
at increasing potash recoveries from high-clay 
ores have been studied by the Carlsbad com- 
panies and by the Bureau of Mines, but none 
of these studies has progressed to the commer- 
cial stage. 

Sylvite ore recovery can also be affected by 
the presence of kieserite (MgS04-H20). If the 
kieserite is fine grained and interlocked with the 
potassium chloride, it may be impossible to lib- 
erate the two minerals except at very fine sizes. 
This fine size will affect the product grade and 
the KjO recovery that can be achieved in the 
flotation circuit. The presence of kieserite in a 
crystallization circuit may cause the formation 
of glaserite [K3Na(S04)2], which will precipitate 
in the lines and hamper recovery of sylvite. 
Maximum SO4 concentration in ore being proc- 
essed in a dissolution circuit should be 3.5 to 4 
percent. 

Carnallite (KCl-MgCl2-6H20)impurities can 



also affect sylvite recovery. In the flotation proc- 
esses, carnallite affects the viscosity of the brine 
and may cause the formation of fine-grained 
potassium chloride. This fine-grained KCl can 
be lost during flotation without careful control 
of the flotation circuit. One company experi- 
encing high carnallite impurities (5 to 6 percent) 
has modified its circuit to include additional 
cleaner steps in fines flotation in order to elim- 
inate this problem. In a crystallization circuit, 
the presence of carnallite can affect the Mg^^ 
ion concentration in the brine leach, causing sig- 
nificant KCl losses. To avoid this, a preleach 
circuit for carnallite is necessary, followed by a 
recrystallization stage to recover the KCl that 
dissolves in the preleach. 

Kainite (MgS04-KCl-3H20) is an impurity in 
sylvite ore that can affect the sylvite product con- 
centrate grade. In flotation, the kainite will also 
float and be recovered as an impurity in the 
concentrate. 

In langbeinite ores, polyhalite 
(K2S04-MgS04-2CaS04-2H20) is an impurity 
that cannot be leached from insoluble langbein- 
ite because of similar solubility characteristics. 
The recovery of polyhalite in the langbeinite 
concentrate will thus be comparable with lang- 
beinite recovery. Required market grade for 
langbeinite is 22 percent KjO with a maximum 
of 4 percent impurities. It is difficult to make 
the required grade if polyhalite impurities are 
high. 

Clay and slimes can also affect langbeinite ores 
if sufficient desliming is not included in the 
process. It is not unusual to expect 20 percent 
of clay and slimes in the feed to report to con- 
centrate. Therefore, control must be maintained 
to lower this amount to meet market grade. 

Theoretically, kieserite should be water leach- 
able and thus not affect the recovery of lang- 
beinite. Plant experience, however, seems to 
indicate that the kieserite does not completely 
leach and may report to the concentrate, thus 
lowering the grade. 

BENCH-SCALE METALLURGICAL TESTS 

As noted previously and shown in table 5, 
many of the potash impurities that affect lang- 
beinite and sylvite recovery are present in the 
WIPP site ores. To understand the effect of the 
impurities on recovery, the Bureau of Mines Salt 
Lake City Research Center conducted prelimi- 
nary sylvite flotation and langbeinite leach tests 
on two core samples taken from drill sites near 
the WIPP area. The primary purpose of these 
tests was to determine if any ore characteristics 
would preclude beneficiation by current flota- 



30 



tion or leaching technology. The tests were de- 
signed to estimate the range of potash grade and 
recovery expected for the ores tested and to 
determine concentrate grades ancPimpurities. 
No attempts were made to reclean concentrates 
or to include secondary ore-desliming steps in 
the metallurgical tests. The addition of such 
procedures could make the concentrates higher 
in grade, with a possible decrease in recoveries, 
but such tests are beyond the scope of this pre- 
liminary study. The tests performed, however, 
could indicate where such recleaning would be 
necessary. The results of these metallurgy tests, 
in conjunction with previously discussed current 
technology and known impurity problems, were 
used to estimate the expected metallurgical 
characteristics of the WIPP site ore. 

The core samples used were AEC-7 and 
AEC-8; these cores were obtained from a pre- 
vious drilling investigation in the area, which 
was reported in a U.S. Geological Survey open- 
file report {10). These core samples were used 
because they provided a larger amount of ore 
for brine makeup and metallurgical tests than 
would have been available by using the 2-inch 
ERDA drill cores from this investigation. Table 
5 summarizes the chemical analyses made by the 
Salt Lake City Research Center on portions of 
the core used in the tests. The table also shows 
a calculated chemical analysis based on the 
uses report for these same core portions. Var- 
iations between the two analyses are due to slight 
differences between core sample splits. Because 
of the close comparison between the Bureau of 
Mines sample analyses and those made by the 



U.S. Geological Survey, the mineralogy ex- 
pected for the Bureau of Mines samples should 
be comparable with that determined in the 
USGS analysis. The author of the USGS open- 
file report (10) on the two core samples, C. L. 
Jones, also supervised the core drilling within 
the WIPP site for this evaluation. Discussion 
with C. L. Jones confirmed that the mineralogy 
shown in the USGS report could be assumed to 
be the mineralogy of the Salt Lake City test ores. 
Further, he indicated that the test ores were rep- 
resentative of ores found in the WIPP area. 
Table 6 shows the assumed mineralogy. 

Sylvite Flotation Tests 

Sylvite flotation tests were performed by the 
Bureau of Mines Salt Lake City Research Center 
on two ore samples, shown as AEC 7-5 and AEC 
8-10 in tables 5 and 6. The laboratory proce- 
dure included the following steps: 

1. The ore at minus 10 mesh was scrubbed for 
15 minutes in a 50-percent solids saturate 
brine. 

2. The scrubbed ore slurry was diluted to 30 
percent solids and then decanted over a 150- 
mesh screen. This desliming operation is sim- 
ilar to the laboratory procedure used in the 
Carlsbad potash industry. The oversize (plus 
150-mesh) fraction was diluted (22 to 33 per- 
cent solids) and deslimed again. The under- 
size (minus 150 mesh) was discarded. 
Commercially, the undersize would be thick- 
ened and discarded and the separated brine 
recycled. 



TABLE 5. — Chemical analysis (percent) of core samples used in Bureau of Mines 

metallurgical test (approximate chemical analysis based on U.S. Geological Survey Open-File 

Report 75—407 included for comparison) 





AEC 7-5 


AEC 8-10 


AEC8-4A 


AEC8-4B 


AEC 8-4C2 


Ore used 


5th ore zone 
AE07 


10th ore zone 
AEC-8 


4th ore zone 
AE08 


4th ore zone 
AE08 




Core 


AEC-8 


Depth, feet _ 


1726.5-1741.5 


1589-1594.5 


1753-1756.5 


1752.0-1756.5 


1753.0-1756.7; 


KjO: 

USBM 










1752-1753 


13.4 


13.1 


12.6 


9.50 


3.70 


USGS 


15.75 


13.4 


12.8 


10.74 


4.56 


Calcium: 












USBM 


.42 


.20 


<.20 


<.20 


<.20 


USGS 


.42 


.21 


.05 


.06 


.07 


Magnesium: 












USBM 


2.10 


3.40 


6.60 


5.27 


1.92 


USGS 


2.80 


3.70 


6.36 


5.00 


2.20 


Chlorine: 












USBM 


44.70 


40.60 


24.38 


31.90 


45.60 


USGS 


44.92 


41.56 


26.41 


33.53 


48.30 


so,: 












USBM 


12.50 


14.90 


39.50 


29.70 


12.15 


USGS 


12.51 


13.70 


38.122 


29.94 


13.05 


Water-insoluble matter: 












USBM 


.80 


4.50 


.70 


1.20 


2.50 


USGS 


.81 


5.53 


1.104 


1.18 


1.33 


Water loss, 60°-200° C: 












USBM 


(') 


(') 


o 


n 


n 


USGS. 


1.34 


1,73 


.64 


.53 


.299 



' Tables used in calculation from USGS open-file report (9). 

' Sample AEC 8-4C is a synthetic sample consisting of 1,500 grams of sample AEC 8-4A (depth 1752.7-1756.7) mixed with 3,200 grams of ore from core AEC-8, 

4th ore zone (depth 1752-1753). 

Water loss analysis not conducted by the Bureau of Mines (USBM). 



31 



TABLE 6. — Approximate mineral content (weight-percent) of core samples used in Bureau 
tests. Mineralogical analysis calculated from data reported in U.S. Geological Survey Open- 
File Report 75-407' 



Mineral 


Chemical 
composition 


Test 
AEC 7-5 


Test 
AEC 8-10 


Test 

AEC8--1A 


Test 
AEC8^B 


Test 
AEC 8-4C2 


Sylvite 


KCI 

K2SO4 ■ 2MgS04 

MgS04 • kA ■ Sh,o 

K2SO4 • MgS04 • 4H2O 

K& ■ MgCl2 • 6H,0 
K^04 ■TvlgS042CaS02 

NAcf 
CaS04 


21.65 
.28 

7.91 




10.52 

3!09 

55.11 
.06 
.83 


20.0 








18.87 

4.20 
.31 

50.2 
.60 
5.52 


0.05 
51.77 
1.04 
2.31 
.34 



43.23 
.09 
1.08 


0.10 
43.14 
1.00 
1.93 
.29 


:3i 

54.88 
.07 
1.16 


29 






Leonite 

Bloedite _ 

Kieserite _.. 

Carnallite 


1.29 
.74 
.11 


.33 


Halite ___ 


78.91 


Water-insoluble matter 


1.31 



Tables used in calculation from USGS open-file report (9). 
Sample AEC 8-4C is a synthetic sample and does not represent a particular 
ics of very low langbeinite. 



t does, however, give an indication of leaching characteris- 



3. The deslimed sample is then conditioned 
prior to flotation. The plus 150-mesh solids 
were diluted to 23 percent solids, and then 
the reagents were added in two steps. Com- 
mercially, conditioning could be done in a 
tumbler at a higher percentage of solids, and 
such conditioning would then use less re- 
agent than laboratory tests because of the 
smaller dilution and possibly greater contact. 

a. In the first stage, a blinder reagent is 
added to the slurry which is then condi- 
tioned for 2 minutes. The blinder keeps 
insoluble slimes from absorbing amine 
and floating. The portion of feed and re- 
agent used are as follows: 

0.2 pound (9 1 grams) MRL 20 1 ' for tests 
with core AEC 7-5 

0.3 pound (136 grams) MRL 201 for 
tests with core AEC 8-10 

b. In the second stage, additional flotation 
reagents were added to the slurry, which 
is then conditioned for 2 minutes. Both 
core samples used identical amounts of 
the following reagents per ton of feed: 

0.2 pound (91 grams) of Armeen T. D., 

an amine collector 

0. 1 pound (45 grams) Barretts oil 634 
0.018 pound (8.17 grams) Hexanol 

frother 

4. The conditioned sample is floated for 2 min- 
utes to produce a rougher concentrate. 

5. The rougher concentrate is then recondi- 
tioned prior to any cleaner flotation. This 
additional conditioning step was found nec- 
essary in the laboratory, possibly due to di- 
lution of the sample, or due to impurities 
present in the ore. Commercially, only initial 
conditioning (step 3) is needed. In condition- 
ing, the rougher concentrate is diluted to 23 
percent solids (for AEC 7-5 sample) and 24 



' Use of specific brands of reagents does n 
of these reagents in preference to simila 



'agents sold by other companies. 



percent solids (for AEC 8-10 sample), and 
is conditioned for 2 minutes with the addition 
of 0. 1 pound (45 grams) of MRL 201 blinder 
per ton of original feed. 

6. The conditioned rougher concentrate is then 
floated for 2.5 minutes to produce a cleaner 
concentrate. 

7. The first cleaner concentrate is then condi- 
tioned for 2 minutes by adding the following 
reagents per ton of original feed: 

0.05 pound (23 grams) Armeen T. D. 
0.024 pound (11 grams) Barretts oil 634 
0.01 pound (4.5 grams) Hexanol 

8. The conditioned cleaner concentrate is then 
floated for 1.5 minutes in a slurry of 10 per- 
cent solids. The concentrates and tailings are 
then analyzed. A discussion of the test results 
for each ore sample follows. 

Sample AEC 7-5— Sylvite (USBM analysis): 
13.4 weight-percent K^O equivalent 

This sylvite ore sample analysis, as shown in 
tables 5 and 6, contains both kainite and kies- 
erite. In addition, the KgO content in this sample 
is slightly lower than that in Carlsbad ores cur- 
rently used for sylvite recovery by flotation. The 
kainite in the sample would be expected to float, 
thus lowering the recovery and concentrate 
grade of the sylvite. The kieserite can also affect 
the concentrate grade, since it is often inter- 
locked with sylvite and cannot be liberated. In 
this sample, however, microscopic examination 
of this sample indicated that kieserite is not en- 
trained in the sylvite, and thus the two might be 
separated. The flotation results are shown in 
table 7. A recovery of 8 1 . 1 percent of the potash 
value was achieved, but the KgO concentrate 
grade was only 55.26 percent rather than the 
market grade of 60 percent. Analysis of the con- 
centrate, as expected, did indicate the presence 
of kainite. Because of this kainite, which could 



32 



not be mechanically separated from the ore, re- 
covery of market grade sylvite by flotation may 
not be possible. Kainite may not affect a solution 
crystallization circuit as detrimentally, and thus 
sylvite may be recoverable from this material by 
solution-crystallization methods. Extensive so- 
lution and crystallization tests were not con- 
ducted due to the limited scope of these 
preliminary tests. 

Sample AEC 8-10— Sylvite sample: 13.1 weight- 
percent K^O equivalent 

This sylvite sample, as shown in tables 5 and 
6, contains kieserite and, in addition, the KjO 
content is slightly lower than those currendy 
used for sylvite flotation recovery. The kieserite 
can affect recovery and the concentrate grade 
of the sylvite since it is often interlocked with 
sylvite and cannot be liberated. However, mi- 
croscopic analysis indicates that kieserite is not 
entrained with the sylvite and should not affect 
flotation. Sample AEC 8-10 has no kainite im- 
purities but has a very large insoluble content. 
The flotation results are shown in table 8. A 
recovery of 73 percent of the potash value was 
achieved, but the KjO concentrate grade was 
only 54. 14 percent rather than the market grade 
of 60 percent. Analysis of the concentrate in- 
dicated the presence of schoenite, with occluded 



halite, fine halite, and insoluble slimes. It might 
be possible to further upgrade this ore by re- 
cleaning to meet market specifications, although 
such recleaning would reduce Motash recovery. 
Sylvite from this type of ore may also be re- 
coverable by solution and crystallization meth- 
ods, where the insolubles can be more effectively 
removed. 



Langbeinite Leach Tests 

Langbeinite leach tests were performed by the 
Bureau of Mines Salt Lake City Research Center 
on three samples, shown as AEC 8-4A, AEC 8- 
4B, and a composite sample AEC 8-4C in tables 
5 and 6. The laboratory procedures were based 
on similar laboratory leach methods used by the 
Carlsbad potash industry. The tests included the 
following steps: 

1. The sample is separated into a plus 14-mesh 
and minus 14-mesh fraction. 

2. The coarse material (minus 4 to plus 14 
mesh) is leached with cold water in an agitator 
for 2 minutes. Leach water requirements 
were calculated to give a 12 percent chloride 
brine if 100 percent of contained halite dis- 
solved during leaching (steps 2 and 4). 

3. The leach liquor is decanted and filtered 
from the insoluble coarse product. 

4. The fine material (minus 4 mesh) is leached 



TABLE 7. — Chemical assay mass balance of sylvite flotation: test on sample AEC 7-5 

(Weight-percem) 



Product 


Percent of 
total input 
tonnage 


K^G 


Insol- 
ubles 


Magne- 
sium 


SO, 


Sodium 


Chlor- 


Cal- 


Feed: 
Coarse 

Slime (discarded) 


85.9 
14.0 


14.06 
9.34 


0.23 
4.28 


2.02 
2.61 


11.90 
16.16 


22.41 
17.36 


45.58 
39.27 


0.38 
.66 






Toul feed 


100.0 
20.5 
19.66 


13.4 
53.01 
55.26 


.80 
.18 
.14 


2.1 

.85 
.76 


12.5 
5.17 
4.76 


21.7 
2.51 
1.55 


44.7 
42.52 
42.64 


.42 


Rougher concentrate 


.08 


Cleaner concentrate 


.07 


Tailings: 
Rougher 


65.5 
.90 


1.87 
4.22 


.25 
.89 


2.38 
2.9 


14.01 
14.0 


28.64 
23.3 


46.54 
39.77 


.47 


Cleaner 


.44 


Total tailings 


66.4 


1.90 


.26 


2.39 


14.01 


28.57 


46.45 


.47 



TABLE 8. — Chemical assay mass balance of sylvite flotation: test on sample AEC 

(Weight-percent) 



8-10 



Product 


Percent of 
total input 
tonnage 


K^O 


Insol- 
ubles 


Magnes- 


SO,, 


Sodium 


Chlor- 


Cal- 


Feed: 
Coarse... 

Slime (discarded) 


80.9 
19.1 


13.38 
11.92 


1.06 
19.06 


3.06 
4.8 


13.23 
21.96 


20.40 
13.61 


43.94 
26.67 


0.12 
.10 








100.0 
19.8 
17.7 


13.1 
49.79 
54.14 


4.5 
1.12 


3.4 
.88 
.61 


14.9 
3.70 
2.50 


19.1 

4.75 
2.76 


40.6 
44.77 
45.04 








Cleaner concentrate 


.04 


Tailings. 

Rougher 

Cleaner... 


61.1 
2.1 


1.58 
13.13 


1.04 
2.37 


3.77 
3.18 


16.32 
13.94 


25.47 
21.56 


43.67 
42.46 


.14 
.19 


Total tailings 


63.3 


1.96 


1.08 


3.74 


16.22 


25.30 


43.56 


.14 



33 



for 1 minute with the unsaturated leach liq- 
uor separated in step 3. 

5. The leach liquor is then decanted and filtered 
from the insoluble fine product. 

The leach liquor and products were analyzed. 

A discussion of the test results for each sample 

follows. 



Sample AEC 8—4A — Langbeinite sample: 12.6 
weight-percent K^O equivalent 

This langbeinite sample analysis, shown in ta- 
bles 5 and 6, has a higher KgO content than 
Carlsbad ores currently used for langbeinite 
leach recovery. The impurities that affect leach- 
ing, polyhalite and other insolubles, occur in 
small quantities. The leach results are shown in 
table 9. A recovery of 91 percent of the potash 
value was achieved, but only a 21.27 percent 
KgO grade was made. This grade is close to 
minimum market requirements of 22 percent 
KjO. Because analysis of the concentrate indi- 
cates that most of the impurities are insoluble 
slimes, a recleaning of the concentrate should 
increase the concentrate grade. In plant prac- 
tice, ores similar to AEC 8-4A should be proc- 
essable with recoveries in the range of 90 to 91 
percent. 



Sample AEC 8-4B — Langbeinite sample: 9.5 
weight-percent K2O equivalent 

This langbeinite sample analysis, as shown in 
tables 5 and 6, has a KgO content comparable 
or slightly higher in grade than Carlsbad ores 
currently used for langbeinite leach recovery. 
The impurities that affect leaching, polyhalite 
and insolubles, occurred in small quantities. The 
leach results are shown in table 10. A recovery 
of 89.8 percent of the potash value was achieved, 
but only a 21.5 percent KgO grade was made. 
Because analysis of the concentrate indicated 
that most impurities are insoluble slimes, a re- 
cleaning of the concentrate should increase the 
concentrate grade. In plant practice, ores similar 
to AEC 8-4B should be processable with recov- 
eries in the range of 88 to 90 percent. 

Sample AEC 8—4C — Langbeinite sample: 3.7 
weight-percent K^O equivalent 

This sample analysis, listed in tables 5 and 6, 
is a composite and does not represent a specific 
deposit found in the area. It does, however, give 
some indication of the possibility of leaching 4- 
percent KjO-equivalent langbeinite. The im- 
purities, polyhalite and insolubles, are in low 
percentage, but the low grade would make any 



TABLE 9. — Chemical assay mass balance of langbeinite leach: test on sample AEC 8-4A 

(Weishl-percent) 



Product 


Percent of 
total input 
tonnage 


K.p 


Insol- 
ubles 


Magnes- 


SO, 


Sodium 


Chlorine 


Feed: 


61.1 
38.9 


13.4 
11.3 


0.43 
1.11 


6.84 
6.22 


42.41 
34.93 


14.88 
16.47 




Fine 


26.66 




100.0 


12.6 


.70 


6.60 


39.50 


15.50 








Concentrate: 

Coarse 

Fine __ 


35.81 
18.16 


21.4 
21.0 


.21 
.97 


11.1 
11.2 


69.0 
68.8 


.12 



.21 



Total concentrate 

Brine solution __ 


53.97 
46.03 


21.27 
2.43 


.47 
.97 


11.13 
1.29 


68.93 
4.99 


.08 
33.58 


.14 
52.80 







TABLE 10. — Chemical assay mass balance of langbeinite leach test on sample AEC 8-4B 

(Weight-percent) 



Produ 

Feed: 
Coarse _ 

Total 

Concentrate: 

Coarse 

Total concentrate 
Brine solution 



Percent of 
total input 



34 



impurities in the concentrate detrimental to 
making market grade. The leach test results are 
shown in table 11. A recovery of 87 percent of 
the potash value was achieved, but only a 19.48 
percent KjO grade was achieved. An analysis of 
the concentrate indicated the presence of im- 
purities; however, more than 80 percent of the 
impurities were already leached in the test. It 
is doubtful if deposits of this grade could be 
beneficiated unless it contained virtually no po- 
lyhalite or insolubles as impurities or unless 
these impurities were concentrated in certain 
size fractions that could be discarded. 

Summary of Metallurgical Tests 

The tests conducted by the Bureau of Mines 
Salt Lake City Research Center confirmed the 
problems expected in ores found in the WIPP 
area. The WIPP ore types seem to behave sim- 
ilarly to ores currently processed in the Carlsbad 
area. High percentages of polyhalite and insol- 
ubles can lower the concentrate grade of lang- 
beinite ores. If the langbeinite ore has a very 
low KjO content, impurities may be large enough 
to require recleaning processes, or the impuri- 
ties may not be removable, resulting in an un- 
marketable product. High impurities of kainite, 
kieserite, and insolubles in sylvite ores can affect 
the KjO recovery and grade of concentrate and, 
in the extreme, can make it difficult to reclean 
concentrates to meet market specifications. 

MINERALIZATION IN THE WIPP SITE 
FROM U.S. GEOLOGICAL SURVEY DATA 

The principal ore zones in the study area are 
the 10th and 4th; subordinate ore zones in the 
area include the 8th, 3d, and 2d. The ore zones 
dip gently to the east-southeast at approximately 
100 feet per mile (about 19 meters per kilo- 
meter). 

The mineralized zones comprise lenticular 
bodies lying between anhydrite and/or thick salt 
beds. Normally, seams of clay and salt occur at 



the top and bottom of these lenses. In several 
places, however, potash mineralization occurs 
above or below the designated ore zone, and 
experience dictates that these occurrences are 
localized and not laterally persistent. This situ- 
ation may exist within the WIPP area; where 
these spotty occurrences are present, they have 
not been included in the ore-reserve estimate 
(figs. 18, 19, and 20). 

The designated ore zones occur as laterally 
persistent beds of halite and argillaceous halite 
that locally contain complex potash minerals in 
various concentrations. The zones are persist- 
ent, but commercial grades within the zones oc- 
cur at irregular intervals. Thus, these persistent 
beds which are, by usage, called ore zones occur 
over large areas, including areas where potash 
mineralization is not present in commercial con- 
centrations. 

The occurrence of potash minerals in the ore 
zones is typically massive, coarsely crystalline, 
fairly even-grained, and mineralogically com- 
plex. The mineralized bodies contain halite and 
one or more potassium minerals as major con- 
stituents, at least one magnesium mineral, and 
small amounts of clay, silt, polyhalite, and/or 
anhydrite. An example of KgO occurrence is il- 
lustrated in a list of minerals and grade by ore 
zone for test hole AEC-8 in table 12. 

Grades and tonnages of potash mineralization 
supplied by the U.S. Geological Survey were 
divided into three basic grade categories, which 
include both measured and indicated amounts. 
The criteria for the three categories were (1) 
highest grade, a minimum thickness of 4 feet 
with a minimum grade of 8 percent K2O as lang- 
beinite, 14 percent KjO as sylvite, or mixed ore 
of comparable grade; (2) medium grade, a min- 
imum thickness of 4 feet (1.2 meters) with a 
minimum grade of 4 percent KjO as langbeinite, 
10 percent K2O as sylvite, or mixed ore of com- 
parable grade; (3) lowest grade, a minimum 
thickness of 4 feet (1.2 meters )with a minimum 
grade of 3 percent KgO as langbeinite, 8 percent 



TABLE 11. — Chemical assay mass balance of langbeinite leach test on sample AEC 8-4C 

(Weight-percent) 



Product 


Percent of 
total input 
tonnage 


K2O 


Insol- 
ubles 


Magnes- 
ium 


SO4 


Sodium 


Chlorine 


Feed: 

Coarse - 

Fine - 


74.8 
25.2 


2.88 
6.18 


2.47 
2.60 


1.49 
3.21 


9.31 
20.59 


32.62 
22.2 


50.06 
32.37 


Total 


100.0 


3.7 


2.5 


1.92 


12.15 


30.0 


45.6 


Concentrate: 

Coarse _ 

Fine 


9.98 
6.53 


19.2 
19.9 


2.09 
2.28 


10.2 
10.1 


66.1 
67.1 


1.08 

.5 


1.37 
.33 


Total concentrate 

Brine solution 


16.51 
83.49 


19.48 
.58 


2.17 
2.57 


10.16 
.29 


66.50 
1.40 


.85 
35.76 


.96 
54.43 



35 



TABLE 12. — Ore zone thickness and grade in test hole AEC-8 

(After C. Jones, 1975) 





Ore 


Depth 

(feet) 


Thickness 

(feet) 


KgO distribution 


by minerals' 






Syl 


Lan 


Lee 


Kai 


Car 


11th... __ 


1.521.8-1,523.1 


1.3 


1.5 








0.5 


10th 




1,589.7-1,594.7 


5.0 


13.6 








1.1 


9th 




1,604.3-1,607.7 


3.4 


3.9 








3.4 


8th 




1.636.6-1,638.1 


1.5 


11.9 








.9 


7ch 




1,666.5-1.671.0 















6th 




1,681.9-1,683.3 















5th 




1,688.7-1,697.0 















4th 




1,753.4-1.757.4 


4.0 




il.o 


0.6 


0.5 




3d 




1,766.0-1,767.0 


1.0 




3.4 


.6 


.5 




2d 




1,781.*-1,782.6 
1,796.0-1,810.5 


.7 













1st 





' Syl = Sylvite; Lan = Langbeinite; Leo = Leonite; Kai = Kainite; Car = Carnallit( 



KjO as sylvite, or mixed ore of comparable 
grade. Mineralized areas, as determined by the 
U.S. Geological Survey and fitting the preceding 



criteria, are shown in figures 18-20. Exploration 
drill hole sample analyses, thicknesses, and 
depths are listed in table 13. 



36 



TABLE 13. 



Calc 


ulat 


ed mineral 


content of selec 


ted 


samp 


les 


trom 


pot 


assium-bea 


ring 


Intervals wi 


th 


summa 


tlon 






of perce 


nt K 


as ore mlr 


era 


1 





Drillhole No.: Drillhole designations; P, E 
FC, Farm Chemical Res. Dev. Corp; IMC, Int 
NFU, Farmers Edu . and Coop. Union of Ameri 
U, U.S. Potash Co. , Inc. 



irch and Development Administ 
Minerals and Chemical Corp; 
il Sulphur and Potash Co. ; 



Calculated Mi 
C, carnallit 
Le, leonite; 



ler headings: Arc. arcanit 
glauberite; Ka , kainite; 
jite, Va, vanthoffite 



Bl, bloedite; 



LUhole Ore 
No. Zone 



Thicknes 
(feet) 



1440.47-1441.35 
1441.35-1442.30 
1442.30-1443.50 



1627.15-1628.35 



1628.35-1629.52 



Calculated minerals pr 
(weight percent) 



Halite Sylvite Langbeinite 



0,88 
0', 9,5 
1.20 



1363.70-1365.00 1.30 
1365.00-1366.72 1.72 
1366.72-1368.05 1.33 



K.O as Weighted average 
ore K^O as ore mineral 

(percent) by ore zone 

(feet and percent) 



5 


1802 


70-1804 


00 


1 


30 




6 


1804 


00-1805 


00 


1 


00 




7 


1805 


00-1805 


85 





85 




8 


1805 


85-1806 


30 





45 




I 


1833 


08-1834 


00 





92 




2 


1834 


00-1834 


50 





50 




3 


1596 


30-1597 


60 


1 


30 




4 


1597 


60-1598 


70 


1 


10 




5 


1598 


70-1599 


53 





83 


34 


2 


1572 


60-1574 


97 


2 


37 




3 


1574 


97-1576 


17 


1 


20 




4 


1576 


17-1577 


77 


1 


60 




5 


1577 


77-1578 


69 





92 




6 


1546 


69-1548 


65 


1 


96 




7 


1548 


65-1549 


66 


1 


01 




8 


1549 


66-1551 


40 


1 


74 




9 


1551 


40-1552 


75 


1 


35 




12 


1476 


00-1477 


45 


1 


45 




13 


1477 


45-1478 


37 





92 




14 


1478 


37-1480 


00 


1 


63 




18 


1510 


50-1511 


32 





82 




19 


1511 


32-1512 


10 





78 




20 


1512 


10-1513 


05 





95 




2 


1479 


73-Ut;i 


20 


1 


47 




3 


1481 


20-1483 


00 


1 


80 




4 


1483 


00-1483 


48 





hS 





76 
86 




15.0 
4.4 


Tr/Ka-' 


3.4/L 
1.0/L 


79 






14.0/Kj 
Tr/Bl-' 












95 




2.34 




0.53/L 


79 




17.07 




3.87/L 


39 




39.9 


1.9/Ka 
5.0/Le 


9.05/L 


46 




43.1 


2,6/Ka 
0.5/Le 


9.78/L 


39 
80 




38.0 


12.9/Ka 


8.6/L 


89 
60 




20.0 


4.0/Ka 
3.7/Le 


4.54/L 


79 





4.4 


1.9/Ka 


3.14/L 


38 


--- 


52.9 


0.5/Ka 


12.0/L 


72 


-_- 


24.7 


0.5/Ka 


5.6/L 


62 




18 , 6 


0.5/Ka 


4,22/L 


38 




23.4 




5.31/L 


56 


46.0 


... 


... 


29.40/S 


64 


27.6 








17.48/S 


86 


7.0 








4.32/S 


64 


28.0 


--- 


... 


17.46/S 


65 


28.0 


... 


... 


18.28/S 


76 
60 


13.3 
21.0 


... 


--- 


8.42/S 
13.41/S 


65 


14.5 


... 


... 


9.14/S 


74 




22.0 


1.0/Ki 


5.01/L 


83 


2.0 


8-0 


6.0/Ki 


1.90/L 
1.26/S 


84 




12.0 




2.78/L 


37 


2 


54.0 


6.0/Ki 


12.31/L 


53 




42.0 




9.65/L 


79 


3.0 


1.0 


8.6/Ka 


0.19/L 


65 


1.0 


",:> 


... 


6.53/L^ 


89 


2.0 





0.75/L 


69 


2.0 


3.8 


5.7/Ka 
1.0/Le 
1.0/Bl 


3.8/L 


52 


29.0 


25.0 




18.45/S 
5.62/L 


46 


3.0 


45.0 


4.0/Ka 
2,0/Le 


10.3/L 
I.65/S 


86 


-_. 


14.0 


.-- 


3.23/L 



2.37-9.41/L 

3.60-3.67/L 
1.42-6.26/L 

3.23-5.06/L 
6.09-18.41/S 
6.06-13.2/S 



2.55-6.98/L 



3.75-3.41/L 



35 
35 


6 
6 


74/L 
17/S 


35 


i equivc 
-9.2/L 



37 



TABLE 13. - Calculated mineral cont 



f selected samples 



iOtassium-bearing intervals with summation 
of percentK.O as ore mineral 



Drillhole No.: Drillhole designations; 
FC, Farm Chemical Res. Dev. Corp; IMC 
NFU, Farmers Edu. and Coop. 1 
U, U.S. Potash Co. , Inc. 



Energy Research and Development Adminis 
nternational Minerals and Chemical Corp; 
f America; D. Duval Sulphur and Potash Co.; 



Calculated Minerals Present 
C, carnallite, Gl . glaserit 
Le, leonite; Lo , loeweite; 



nd other headings: Arc, arc 
; Gu , glauberite ; Ka , kaini 
, sylvite , Va , vanthoffite 



Lte; Bl, bio 
L, langbein 



Calculated 



Drillhole Ore Sample 
No . Zone No . 



Thickness 
(feet) 



Polyhalite Halite Sylvite Langbeinit 



K as Weighted average 
ore K as ore mineral 
Other minerals for intervals preceding 
minerals (percent) by ore zone 

(feet and percent) 



1522.56-1523.41 



1523.41-1524.07 
1524.07-1524.70 



11 1524. 70-1526. 0« 

12 1526.04-1526.74 



1703, 


,65- 


■1705, 


,23 


1.58 


1705 


.23- 


-1705, 


,65 


0.42 


1650 


,38- 


-1651, 


,22 


0.84 


1651, 


.22- 


-1652 


,03 


0.81 


1562 


,03- 


-1653 


,83 


1.80 


1653, 


.83- 


-1654 


,58 


0.75 


1601, 


,90- 


-1603 


,56 


1.66 


1603, 


,56- 


-1604, 


,64 


1.08 


1604, 


,64- 


-1605 


38 


0.74 



11 4 

10 7 1670.70-1671.84 

10 8 1671.84-1673.42 

10 9 1673.42-1674.70 



1868.67-1870.28 1.61 
1870.28-1871,10 0.82 
1871,10-1872.30 1.20 



71 19.0 
66 10.0 
71 16.7 



1688.72-1689.60 


0.88 


3 


76 


22.0 


1689.60-1690.89 


1.29 


1 


64 


36.9 


1690.89-1691.95 


1.06 


1 


71 


21.3 


1691.95-1693.28 


1.33 


2 


92 


2.0 


1840.60-1842.35 


1.75 


1 


58 


___ 


1842.35-1843.40 


1.05 


4 


76 


5.0 



48.0 


0.67/Ka 


4.73/S 
10.8/L 




--_ 


3.0/Ki 


0.96/S 




63.37 




4.42/S 
14.37/L 




1.10 





0.24/L 




31.31 




7.1/L 


4.18-5.63/L 

2.14-3.48/S 

mixed ore equivalent 

4.18-6.34/L 


28.70 




6.51/L 




6.80 


3.0/Ka 


1.55/L 


2.00-5.47/L 




1.0/Ka 


1.13/S 






2.0/Ki 


3.83/S 






1.0/Ki 


4.85/S 


3.45-3.70/S 




3.60/Ki 


3.34/S 






1.0/C 






--- 


31.0/Ka 
3.0/C 


1.48/S 






7.0/Ka 


2.20/S 






2.0/C 




3.48-2. 52/S 


8.0 


1.0/Ki 


12.0/S 
1.82/L 




18.0 


1.0/Ki 


6.32/S 
3.98/L 






1.0/Ki 


10.53/S 

13.74/S 
23.30/S 


4.00-9.29/S 

2.72-3.07/L 

mixed ore equivalent; 

4.00-14.51/5 





2.0/C 


13.43/S 




--- 


1.0/C 


1.24/S 


4.56-12.73/S 


40.0 


5.0/Ki 


9.14/L 




4.9 




1.11/L 








2.98/S 


2.80-6.0/L 


60.0 


2.0/Ki 


13.65/L 




45.0 


1.0/Ki 


13.53/L 




58.3 


8.0/Ka 


,13.24/L 


3.63-13.49/L 




Tr/Bl&Lo-' 





Data provided by U.S. Geological Survey. 



38 



Calculated mineral content of selected samples 
from potassium-bearing Intervals with summation 
of percent K.O as ore mineral 



Drillhole No.: Drillhole designations; P, Energy Research and Development Administrati 
FC, Farm Chemical Res, Dev. Corp; IMC, International Minerals and Chemical Corp; 
NFU. Farmers Edu . and Coop. Union of America. D, Duval Sulphur and Potash Co.; 
U, U.S. Potash Co. . Inc. 

:alculated Minerals Present and other headings: Arc. arcanite; Bs , bischofite: Bl , blc 
C. carnallite. Gl , glaserite; Gu , glauberite ; Ka , kainite; Ki , kieserite; L, langbeir 
Le, leonite; Lo , loeweite; S, sylvite, Va , vanthoffite 



K^O as Weighted average 
Drillhole Ore Sample Depths of Thickness ore K.O as ore mineral 

No. Zone No. interval (feet) Other minerals for intervals preceding 

(feet) Polyhalite Halite Sylvite Langbeinite minerals (percent) by ore zone 



(feet and percent) 






lly 


1344.97-1345.27 


0.3 


4 


36 


62.3 





1345.27-1345.90 


0.63 





10 


24.0- 





3 


1345.90-1346.95 


1.05 


1 


59 


18.0 





4 


1346.95-1348.90 


1.95 


3 


64 


18.0 





5 


1348.90-1349.91 


1.01 


7 


44 


22.0 





6 


1349.91-1350.80 


0.89 


3 


67 


17.0 



1392 


66 


1394 


29 


1.63 


1394 


29 


1394 


90 


0.61 


1520 


00 


1521 


55 


1.55 


1521 


55 


1522 


39 


0.84 


1533 


50 


1535 


05 


1.55 


1535 


05 


1535 


59 


0.54 



1.0/Ki 39.38/S 

59.0/Ki 15.08/S 

18.0/Ki 11.39/S 

19.0/Ki 11.36/S 

17.0/Ki 14,06/S 

1.0/Ki 10.69/S 5.83-13.57/S 

9.77/S 
4.12/S 
6.47/S 
22.42/S 4.71-8.24/S 



23.0 45.0 9.0/Ki 14.47/S 

10.13/L 

7.0 6.0 16.0/Ka 1.44/L 



49.0 --- 7.0/S 

8.59/L 

8.0 1.0/Ki 1.80/L 



0.62/S 3.51-5.98/L 

1535.59-1537.01 1.42 2 47 10.0 21.0 --- 4.72/L 3.51-5.74/S 

6.32/S 



1549.79-1550.65 


0.86 


1550.65-1551.29 


0.64 


1551.29-1551.61 


0.32 



5.0 



2.42/L 0.54/L 
l.O/I^^ 
Tr/Le- 15.93/L 



21 5.0 30.0 18.0/Ka 
19.0/Le 

17.0/Bl 12.48/L 1.82-8.05/L 

1 46 49.5 --- --- 31.32/S 

1 62 17.0 2.0 15.0/Ki 10.92/S 

3 97 

1 30 20.0 10.0 30.0/Ki 12.6/S 3.85-14.67/S 
10.0/Ka 2.27/L 

2 82 2.0 9.0 3.0/Ki 2.03/L 
1 63 --- 33.2 0.4/Le 7.53/L 3.53-4.06/L 

1 64 --- 7.0 7.0/Bl 1.59/L 

3 8 1493.23-1493.88 0.65 2 86 --- 7,0 --- 1.59/L 

3 9 1493.88-1494.58 0.70 --- 15 --- 53.0 --- 12.06/L 

3 86 --- 10.0 --- 2.21/L 2.27-5.07/L 

4.00-2.88/L-' 

2 67 30.2 --- --- 19.13/S 

8 80 9.14 --- --- 4.97/S 2.05-11.74/S 



29 


1318 


02-1319 


00 


0.98 


30 


1319 


00-1320 


22 


1. 22 


3^^^/ 


1320 


22-1320 


88 


0.66 


1320 


88-1321 


87 


0.99 


2 


1480 


20-1481 


73 


1.53 


3 


1481 


73-1482 


78 


1.05 


4 


1482 


78-1483 


73 


0.95 


8 


1493 


23-1493 


88 


0.65 


9 


1493 


88-1494 


58 


0.70 


10 


1494 


58-1495 


50 


0.92 


21 


1359 


65-1360 


63 


0.98 


22 


1360 


63-1361 


70 


1 07 



39 



TABLE 13. - Calculated mineral content of selected samples 
from potassium-bearing intervals with summation 



jf percent K„0 aE 



Drillhole No.: Drillhole designations; P. Energy Research and Development Admini 
FC, Farm Chemical Res. Dev. Corp; IMC. International Minerals and Chemical Corp 
NFU, Farmers Edu. and Coop. Union of America; D. Duval Sulphur and Potash Co.; 
U, U.S. Potash Co. , Inc. 



Cal 



ated Minerals Present and other headings: Arc, arcari 
rnallite; Gl, glaserite; Gu, glauberite; Ka , kainite 
;onite; Lo, loeweite; S, sylvite, Va . vanthoffite 



:hofite; Bl . bloedite; 
Lte; L, langbeinite; 



Drillhole Ore 



Depths of 
interval 
(feet) 



Calculated minerals pre 
(weight percent) 



Halite Sylvi 



Weighted average 
K.O as ore mineral 
for intervals preceding 
by ore zone 
(feet and percent) 



9 


37 


1336.38-1337.32 


9 


38 


1337.32-1338.64 





2 


1255.24-1255.64 





3 


1255.64-1257.56 





4 


1257.56-1259.07 





5 


1259.07-1260.15 



L260. 15-1261, 



1440.79-1441. 



19 1441.98-1442.84 

20 1442.84-1443.98 



1443.98-1444.61 



1364.44-1366. 



1366.11-1367.86 
1367.86-1369.26 



1.56 
0.94 



1476 


76- 


1478 40 


1 64 


1478 


40- 


1478.95 


0.55 


1490 


12- 


1491.00 


, 88 


1491 


00- 


1491.56 


0.56 


1491 


56- 


1492.64 


1 08 


1402 


54- 


1493.89 


1.25 



1371.94-1372 


81 





87 


6 


76 


1372.81-1374 


77 


1 


96 




89 


1374.77-1375 


80 


1 


03 


9 


64 


1399.66-1400 


33 





72 


3 


64 


1400.38-1401 


51 


1 


13 




78 


1301.94-1302 


57 





63 


3 


93 


1302.57-1303 


91 


1 


34 


3 


79 


1303.91-1304 


39 





48 


--_ 


61 



--- 


--- 


1,12/S 
21.99/S 




--- 


— 


2.41/S 


3.82-6,7/S 


... 


— 


36.03/S 










36.06/S 






Tr/Kai/ 


8.84/S 




18.0 


3.99/L 




22.0 




4.36/S 
4,88/L 






16.0/Ka 


6.21/S 


5.81-18.47/S 
2.59-4.36/L 
mixed ore equivalent 
5,81-23.33/S 


38.5 


— 


8.74/L 
3.99/S 




23.0 


Tr/Le 


5.22/L 






2.0/Ka 


2.53/S 




11.0 


... 


2.50/L 


3,82-4,69/L 
3.19-2,39/S 


15.0 


1.0/Ka 


3.35/L 
3.00/S 




19.0 


Tr/KaY 


4,3/L 




13.0 


Tr/Kai' 


3.03/L 


4,82-3.60/L 
1.67-3.0/S 
mixed ore equivalent 
4.82-4. 02/L 


25,0 




5.59/L 




4.0 


.-. 


1.00/L 




28.0 


-- 


6.30/L 


3,86-3,45/L 


32,0 


4.0/Ka 


7.18/L 




17.0 


3.0/Ka 








3,0/Le 


4.28/L 


1,85-5,41/L 


4.0 




0.47/S 
0.89/L 




21.0 




4.77/L 




34.0 


8.1/Ka 
8.0/Le 
Tr/Kii' 


7.77/L 


2.45-4.63/L 








53.0 


Tr/Ka&Le-' 


12 01/L 
1.26/S 




8.7 




1.98/L 




S:;i' 





4.94/L 






4.40/L 








4.12/S 




12.0 




2.75/L 




31.0 


6.0/Le 


7.05/L 






2.0/Ka 


2.09/S 


5.96-6.61/S 



Data provided by U.S. Geological Survey. 



40 



TABLE 13. - Calculated mineral content of selected samples 
from potassium-bearing intervals with summation 



)f percent K as ore mineral 



Drillhole No.: Drillhole designations; P, Energy Research and Development Adminis 
FC, Farm Chemical Res. Dev. Corp; IMC, International Minerals and Chemical Corp; 
NFU, Farmers Edu. and Coop. Union of America; D, Duval Sulphur and Potash Co.; 
U, U.S. Potash Co. , Inc. 



Ca 



ed Minerals Presen 
C, carnallite; Gl , glaser 
Le , leonite; Lo . loeweite 



and other headings: Arc, arcanite, Bs , bischofite; Bl , bloe 
5; Gu. glauberite; Ka , kainlte; Ki , kieserite; L, langbeini 
3. sylvite. Va , vanthoffite 



.Ihole Ore Sample Depths of 
). Zone No. interval 



Thicknes 
(feet) 



Weighted average 
K.O as ore mineral 
Dr intervals preceding 
by ore zone 
(feet and percent) 



2 


19 


1526 


90-1528.00 


1.10 





2 


1365 


60-1367.20 


1.6 





3 


1367 


20-1368.45 


1.25 





'' 


1368 


45-1369.70 


1.25 


4 


8 


1542 


90-1543.68 


0.78 


4 


9 


1543 


68-1544.46 


0.78 


2 


18 


1591 


39-1592.71 


1.32 


2 


19 


1592 


71-1594.21 


1.50 



2.0/Ka 9.05/L 



1728.40-1728.78 

1728.78-1730.45 
1730.45-1731.49 
1731.49-1732.29 






68.8 





15 


61/L 





12.6 


2.0/Ka 


2 


87/L 


_-_ 


6.64 


30.0/Le 
9.0/Ka 


' 


51/L 


... 


38.0 


13.0/Le 
Tr/Kai' 


8 


62/L 










... 


53.0 


3.0/Le 


12 


03/L 




46.0 


1.0/Ka 


10 


46/L 




2.4 


2.0/Ka 





53/L 


3.0 


35.0 




22 

7 


83/S 
94/L 


9.0 






5 


61/S 


0.72 











45/S 


0.81 


--- 


_._ 





51/S 



4.10-7.43/L 



. 56-10. 33/L 



3.89-4.86/S 
0.38-7. 94/L 

3.89-6.8/S 



7 


1741.80-1742.35 


0.55 


8 


1742.35-1743.72 


1.37 


9 


1743.72-1745.09 


1.37 


10 


1745.09-1746.00 


0.91 


21 


1925.20-1925.90 


0.70 


22 


1925.90-1926.70 


0.80 


23 


1926.70-1927.94 


1.24 



1956.40-1957.36 


0.96 




1957.36-1958.71 


1.35 




1958.71-1959.21 


0.50 


0.5 


1725.00-1726.15 


1.15 




1726.15-1728.10 


1.95 




1728.10-1729.62 


1.52 




1729.62-1731.48 


1.86 





398.80-1900.45 


1.65 


900.45-1901.77 


1.32 


901.77-1903.35 


1.58 



36.0 


... 


4.23/S 
8.24/L 


59.0 


... 


1.20/S 
13.39/L 


15.0 


... 


0.62/S 
3.40/L 


30.0 


8.4/Le 
2.0/Ka 


9.21/L 


11.0 


12.0/Ki 


2.50/L 


57-P 


2.0/Ka 
9.0/Ki 


12.93/L 


3.6 


5.0/l^a 
Tr/Le-' 


0.82/L 







16.0 15.0/Ki 3.74/L 

65.0 Tr/Kai' 14.83/L 

25.0 Tr/Ka-' 5.72/L 



2.74-4.79/L 



2.81-9.42/L 



4.42/S 




8.01/L 




17.69/S 


1.95-8.01/L 


21.48/S 


6.51-14.03/S 




mixed ore equi 




6.51-20.03/S 


8.35/L 




0.30/L 




4.0/L 


4.55-4.5/L 



41 



Calculated mineral content of selected samples 
from potassium-bearing intervals with summation 
of percent K as ore mineral 



Drillhole No.: Drillhole designations; P. Energy Research and Development Admini 
FC. Farm Chemical Res. Dev . Corp; IMC, International Minerals and Chemical Corp 
NFU, Farmers Edu. and Coop. Union of America; D, Duval Sulphur and Potash Co.; 
U, U.S. Potash Co. , Inc. 



Calculated Minerals Present 
C, carnallite; Gl, glaseri 
Le , leonite; Lo , loeweite; 



id other headings: Arc, 
; Gu, glauberite; Ka , ka 
, sylvite, Va , vanthoffi 



schofi 



Ki, kieseri 
, anhydrite 



; Bl, bloedit 
langbeinite; 



Depths of 


Thickness 


interval 


(feet) 


(feet) 





Halite Sylvite Langbeinit 



K.O as Weighted average 
ore K as ore mineral 
Other minerals for intervals preceding 
minerals (percent) by ore zone 

(feet and percent) 



2 


14-' 


1925 


08-1926 


30 


1.22 


2 


15 


1926 


30-1927 


75 


1.45 


10 


2 


1644 


03-1644 


84 


0.81 


10 


3 


1644 


84-1646 


00 


1.16 


10 


4 


1646 


00-1646 


33 


0.33 


10 


5 


1646 


33-1647 


20 


0.87 


10 


6 


1647 


20-1648 


22 


1.02 


10 


7 


1648 


22-1649 


23 


1.01 


8 


15 


1685 


17-1686 


48 


1.31 


8 


16 


1686 


48-1687 


20 


0.72 


8 


17 


1687 


20-1688 


24 


1.04 


8 


18 


1688 


24-1688 


77 


0.53 


8 


19 


1688 


77-1690 


19 


1.42 


8 


20 


1690 


19-1691 


26 


1.07 


8 


21 


1691 


26-1692 


40 


1.14 


8 


22 


1692 


40-1693 


34 


0.94 


4 


24 


1809 


90-1811 


50 


1.60 


4 


25 


1811 


50-1811 


82 


0.32 


4 


30 


1815 


51-1816 


10 


0.59 


4 


31 


1816 


10-1817 


25 


1.15 



1589.10-1589.70 



1589.70-1591.70 
1591.70-1592 20 



1592.20-1594.50 



1594.50-1594.70 



1594.70-1595.50 



1752.70-1753.40 
1753.40-1754.00 
1754.00-1755.00 
1755.00-1756.70 



2.00 
0.50 



95 






0.88 


... 





2/L 






61 






35.5 


... 


8 


05/L 


2 


67-4. 46/L 


80 


17 









10 


82/S 






81 


11 


26/ 

$ 




2.0/Ki 


7 


46/S 






28 


20 




42.0/Ki 


12 


76/S 






41 


25 





18.0/Ki 


16 


14/S 






62 


29 





1.0/Ki 


18 


66/S 






64 


31 


3 


-- 


— 


19 


81/S 


5 


20-14. 37/S 


55 


45 


> 


... 


... 


28 


4/S 






83 


6 








3 


72/S 






92 


3 











1 


95/S 






95 





8 











52/S 






86 


9 


9 








6 


24/S 






90 


1 














75/S 






65 


33 











21 


16/S 






65 


13 





... 


... 


8 


18/S 


8 


17-10. 24/S 


1.0 


3 



0^ 


64.0 


2.5/Ka 
3.9/Le 


14 


60/L 


1 


92-14. 69/L 
00-7.05/L-' 


18 


5 


54.66 


9.0/Ka 


15 


12/L 


4 










10.60/Le 










51 




5 


27.0 


9.0/Ki 


6 


08/L 






42 




5 


44.0 


8.0/Ka 


9 


95/L 


1 


74-8. 64/L 


85 




5 




5/Ki 
2/An 


2 


97/S 






43 




16 




40/Ki 


10 


32/S ' 






37 




24 




6/C 
30/Ki 
1/An 


15 


16/S 






49 




32 


-r- 


7/C 

9/Ka 

0.8/ An 


20 


31/S 






38 




4 




51/C 
9/Ka 
1/Ki 


2 


34/S 














0.1/An 








85 




3 




4/C 
2/Ki 
9/Ka 
0.9/An 




18/S 


6 


4-12. 33/S 


95 




.. 


4 


... 





86/L 






69 




2 


33 


3/Ka 


7 


39/L 






33 




2 


68 


2/Ka 


13 


90/L 






24 




-- 


69 


3/Le 


15 


38/L 


4 


11.27/L 



42 



Calculated mineral 



;lected samples 



potassium-bearing intervals with summation 
of percent K as ore mineral 



Drillhole No.: Drillhole designations; P. Energy Research and Development Administration; 
FC, Farm Chemical Res. Dev. Corp; IMC, International Minerals and Chemical Corp; 
NFU . Farmers Edu. and Coop. Union of America; D, Duval Sulphur and Potash Co.; 
U, U.S. Potash Co. , Inc. 



Calculated Minerals Present and other headings: Arc, arcanit 
C, carnallice; Gl , glaserite; Gu, glauberite; Ka , kainite; 
Le, leonlte; Lo , loeweite; S, sylvite, Va , vanthoffite 



Bs, bischofite 
, kieserite; L, 



Bl, bloedi 
Langbeinite 



Drillhole Ore Sample Depths of 
No. Zone No. interval 
(feet) 



Polyhalite Hal 



1377.67-1379.00 1.33 
1379.00-1381.00 2.00 
1381.00-1382.00 1.00 





r,." 


Weighted average 




K as ore mineral 
for intervals preceding 


Other 


minerals 


minerals 


(percent 


by ore zone 
(feet and percent) 


20.0/Ki 


7.11/S 




1.0/Ki 


24.72/S 




3.0/C 


12.29/S 


4.33-16.4A/S 


3.0/Ki 







1391.50-1393.00 



1415.08-1415.75 
1415.75-1416.50 
1416.50-1418.00 

1418.00-1419.50 
1419.50-1420.42 

1420.42-1421.50 

1466.00-1467.58 



0.67 
0.75 
1.50 



1467 


58 


1469.00 


1.42 


1469 


00 


1469.67 


0.67 


15?9 


92 


1531.42 


1.50 


1531 


42 


1532 42 


1.00 


1564 


17 


1564.92-' 


0.75 


1564 


92 


1566.13 


1.21 


1566 


13 


1567.38 


1.25 



70 


25 


89 


2 


97 


1 



... 


Tr/Ki^ 
Tr/Kii' 


16 


10/S 





1 


35/S 


... 


1.0/lfa 
Tr/Kii' 











5//S 




1.0/I^a 


' 


62/S 






10 


04/S 


... 


1.0/Ka 


10 


21/S 


19.6 




4 


43/L 


40.7 





9 


22/L 


24.2 


... 


5 


49/L 


52.6, , 
33. 0^' 


... 


11 


93/L 




7 


42/L 


... 


2.0/C 


23 


16/S 


... 


4.0/C 


9 


71/S 


6.1 


10.0/C 


5 


62/S 
38/L 



3.67-6.48/L 



2.5-10.13/L 



3.21-11.26/S 
3.21-0.54/L 
Lxed ore equivale 
3.21-12.61/S 



1687 


00- 


1688 


21 


1.21 


1688 


21- 


1689 


46 


1.25 


1712 


88 


1714 


46 


1.58 


1714 


46 


1715 


46 


1.00 


1715 


63 


1716 


00 


0.38 


1541 


42 


1541 


79 


0.38 


1541 


79 


1542 


29 


0.42 


1542 


29 


1543 


96 


1.67 



1543.96-1544.50 
1613.42-1615.08 
1615.08-1615.92 



74.0 


4.0/lfa 
Tr/Vag 
Tr/Va^' 


16 


71 








67.0 


15 


31/L 


56.0 


2.2/Le 


12 


7/L 


30.3 


3.0/I^a 


6 


9/L 








42.8 


Tr/Le^/ 


9 


71/L 


72.2 


2.0/Le 








6.0/C 


16 


34/L 


7.2 





I 


63/L 


49.8 


4.0/Lo 
20.9/C 
Tr/Lei' 


11 


3/L 


20.0 


'* 


54/L 


63.0 


9.3/l^i 
Tr/Loi' 








14 


3/L 


12.8 


17.7/Le 
12 0/Ki 


2 


9/L 



2.46-16.0/L 



3.96-10.4/L 



2.5-10.52/L 



43 



314-720 



TABLE 13. - Calculated mineral content of selected samples 
from potassium-bearing Intervals with summation 
of percent K.O as ore mineral 



drillhole No.: Drillhole des 
FC, Farm Chemical Res. Dev. 
NFU, Farmers Edu. and Coop, 
U, U.S. Potash Co. , Inc. 



Lgnations; P. Energy Research and Development Admin 
Corp-. IMC, International Minerals and Chemical Cor 
Union of America; D, Duval Sulphur and Potash Co.; 



alated Minerals Present and other headings: Arc, arcanit 
:arnallite; Gl . glaserite; Gu , glauberite; Ka. kainite; 
leonite; Lo , loeweite; S, sylvite; Va . vanthoffite 



bloedite; 

beinite; 



Drillhole Ore Sample 
No . Zone No . 



Calculated minerals present 
(weight percent) 

K^O as 
ore ^2" 
Other minerals for intervals preceding 
Polyhalite Halite Sylvite Langbeinite minerals (percent) by ore zone 

(feet and percent) 



Weighted average 
K.O as ore mineral 



1624.33-1625.58 



1625.58-1627.00 


1.42 


1627.00-1627.75 


0.75 



L712. 25-1712. 66 



L712. 66-1713. 



18 


1715.75-1715.75 


1.00 


19 


1715 .75-1716.33 


0.58 


tO 


1718.66-1719.42 


0.75 


a 


1719.42-1720.75 


1.33 


i2 


1720.75-1722.00 


1.25 



43 1722.00-1722.58 0.58 

44 1742.75-1743.25 0.50 

45 1743.25-1744.25 1.00 

46 1744.25-1745.00 0.75 



1604.92-1606.04 1.12 
1606.04-1606.50 0.46 
1606.50-1606.75 0.25 



57 


1740.66-1741.58 


0.92 


58 


1741.58-1742.25 


0.66 


59 


1742.25-1743.25 


1.00 


60 


1743.25-1744.08 


0.83 


61 


1744.08-1745.00 


0.92 


62 


1769.42-1770.08 


0.66 


63 


1770.08-1771.42 


1.33 



J 65 1772,08-1772.58 5 

n.it.i orovic'ed b\- r.S. leolooical Survey. 



31.1 


3.5/Bl 
3.7/Ka 


7.06/L 






29.9 


0.6/{ 
Tr/Bli' 


6.78/L 






62.2 


14.H/L 








5.0/Gl 




3.42- 


-8.49/L 


50.0 


11.0/Ka 
7.5/Le 
6.0/Ki 


11.35/L 








5.0/Ka 
3.4/Ki 


... 






64.6 


5.0/Ki 


14.66/L 






58.7 


5.0/Ka 
1.3/Ki 


13.32 






47.0 


5.0/l^a 
Tr/Gui' 


10.66 


4. 08- 


■11.3/L 










54.0 


6.2/Ka 
4.0/Gl 
1.2/Ki 


12.25/L 






8.0 


3.0/Ka 
9.0/Bs 


1.82/L 






59.5 





13.5/L 






37.08 


Tr/Gu 
3.4/Bs 


8.41/L 






13.5 


0.4/Ka 
6.0/Bs 
Tr/Gu-' 


3.06/L 


3.91- 


-8.08/L 










62.0 


8.8/|^ 
Tr/Va-' 


14.07/L 














30.5 


10.0/Ki 
10.0/Va 


6.92/L 






1.0 


3.2/Ka 
10.1/Ki 
24.0/ya 
Tr/Gui' 


0.73/L 


2.25- 


•6.44/L 










Si; 


Tr/Ci; 


29.0/S 






Tr/ci' 


31.0/S 






1 


5.7/C 


18.5/S 


1.83- 


-28/S 


46.7 




10.6/L 






61.0 




13.85/L 






61,2 


0.1/Ki 


13.89/L 






61.3 




13.91/L 






48.8 


1.8/Ki 


11.07/L 


4.33- 


■12.6/L 


55.4 


3.0/Va 


12.57/L 






45.7 


2.0/Va 
1.0/Arc 


10.37/L 






33.0 


17.0/Va 
8.0/Ki 
1.0/Gu 


7.49/L 






1,5 


14.0/Va 


0.34/L 


3.15- 


■8.64/L 




1,0/Gu 









44 



TABLE 13. - Calculated mineral conte 



of selected samples 



from potassium-bearing intervals with summation 
of percent K„0 as ore mineral 



Drillhole No.: Drillhole designations; P, Energy Research and Development Adminis 
FC. Farm Chemical Res. Dev. Corp; IMC, International Minerals and Chemical Corp; 
NFU, Farmers Edu. and Coop, Union of America; D, Duval Sulphur and Potash Co.; 
U, U.S. Potash Co. , Inc. 



Calc 



Pre 



headings : Arc , 
auberite; Ka . ke 
3, Va, vanthoffi 



:hofite; Bl , bloedi 
Lte; L, langbeinite 



d minerals present 
ight percent) 



Halite Sylvite Langbe 






ther 
nerals 


^re 
minerals 
(percent) 


K as ore mineral 
for intervals preceding 
by ore zone 
(feet and percent) 





1 


7/Ki 
3/Ka 
35/Le 


0.46/L 
0.44/S 
1.46/L 
0.63/S 


2.4-0.88/L 
2.4-0. 51/S 


2 

2 
1 


0/Ka 
0/Le 
0/Ka 


1.50/L 
3.0/L 
1.3/S 


2.5-2.46/L 
1.5-1.3/S 






24.3/S 
0.1/S 
0.26/S 
7.1/S 


4.0-7,52/S 


1 


0/Le 


0.63/L 
11.65/L 
0.25/L 


3.7-6.5/L 
(4.0-6.03/L)-' 




— 


19.67/S 


3.92-19.67/S 




... 


11.96/L 


2.33-11.96/L 






9.1/S 


4.0-9.1/S 






13.1/L 


4.2-13.1/L 




... 


0.8/S 


4.08-0.8/S 




;;; 


31.63/S 
10.4/S 
7.34/S 


4,2-15.3/S 






13.8/S 


2.2-10.7/S 




... 


0.5/S 


2.2-7.5/L 
mixed ore equivalent: 
2.2-29.45/S 








4.0-11/S 
4.0-2.1/L 
(visual estimate) 

9.26-17.37/S 








2.0-8.8/L 



1430.70-1432.10 


1 


40 


6 


1432.10-1433.10 


1 


00 


13 


1627.90-1629.00 
1629.00-1630.50 


1 
1 


10 
50 


2 
4 


1427.00-1427.70 
1427.70-1428.30 
1428.30-1429.20 
1429.20-1431.00 






1 


7 
6 
9 


1 
2 



1528.00-1528.90 
1528.90-1530.90 
1530.90-1531.70 

1536.25-1540.17 

1646.75-1649.08 

1479.80-1483.80 

1598.50-1602.70 

1441.08-1445.17 

1248.30-1249.30 
1249.80-1251.10 
1251.10-1252.50 

1419.60-1421.30 

1421.30-1421.80 



10 10/ 1236.60-1240. 



10/ 1280.52-1289.78 
10/ 1414.20-1416.20 



3.92 

2.33 

4.00 

4.20 

4.08 

1.50 
1-30 
1.40 

1.70 

0.50 



9.26 
2.00 



provided by U.S. Geological Su 



45 



TABLE 13. - Calculated mineral content of selected samples 
from potassium-bearing Intervals with summation 
of percent K as ore mineral 



rillhole No. : Drillhole des 
FC, Farm Chemical Res. Dev. 
NFU, Farmers Edu . and Coop. 
U, U.S. Potash Co. , Inc. 



.gnati 
Corp; 
Unior 



)ns , P, Energy Research and Development Adminis 
IMC, International Minerals and Chemical Corp; 
of America; D, Duval Sulphur and Potash Co.; 



Calculated Minerals Present and other headings: Arc, ai 
C, carnallite; Gl, glaserite; Gu , glauberite; Ka , kair 
Le , leonite; Lo , loeweite; S, sylvite; Va , vanthoffite 



Bl, bloedit 
langbeinite; 



Calculated minerals prese 



Drillho 


Le Ore 


Sample 


Depths of 


Thlckne 


No. 


Zone 


No. 


interval 
(feet) 


(feet) 



K as Weighted 

ore K^O as or 

Other minerals for inter 

minerals (percent) by or 

(feet 



s preceding 
percent) 



.33 


1533.10-1533.50 


0.40 


34 


1533.50-1533.80 


0.30 


35 


1533.80-1534.40 


0.60 


36 


1534.40-1534.70 


0.30 


37 


1534.70-1535.00 


0.30 


38 


1535.00-1539.80 


4.80 



1.51/1, 
2.43/L 
0.50/L 
0.85/L 
0.14/L 
0.85/L 



-0.90/L 
-22.74/S 
ore equivalent : 



10 
10 


24 
25 


1358.60-1359.60 
1359.60-1360.20 


0. 


10 
10 


> 


1360.20-1360.60 
1360.60-1361.20 


0. 
0, 


10 
10 


28 
29 


1361.20-1361.80 
1361.80-1362.30 


0. 
0. 


10 


.30 


1362.30-1363.10 


0. 



47 



1545.90-1546.30 
1546.30-1546.66 
1546.66-1548.00 
1548.00-1548.60 



52 1552.20-1553.00 

53 1553,00-1553.20 

54 1553.20-1553.60 

55 1553.60-1554.00 

56 1554.00-1554.80 



1.40 
0.60 



32.0 
73.2 



2/Ka 


3.34/S 




2/Ka 


3.34/S 
0.25/L 




6/Ka 


0.31/L 




5/Ka 


3.79/S 
3.70/L 




2/Ka 


2.04/L 




7/Ka 


3.34/S 
2.31/L 




1/C 


2.66/S 


4.5-2.54/S 


3/Ka 




2.7-1.8/L 

mixed ore eq. 

4.5-5. 24/S 


0.74/Le 


7.2/L 




9.0/Ki 


16.6/L 




3.8/Le 


6.89/L 




--- 


12.4/L 


2.7-9.24/L 




1.05/L 




2.0/Le 


9.41/L 
3.59/L 




1.32/Le 


11.19/L 
2.0/L 






0.9/S 


2.6-3.94/L, 



1340.40-1341.70 
1341.70-1343.20 
1343.20-1344.00 
1344.00-1344.70 
1344.70-1346.40 



1.30 
1.50 
0.80 
0.70 
1.70 



6.0-5.5/S 
3.2-4.71/L 



6.0-11.78/S 



1391.10-1392.00 
1392.00-1395.30 
1395.30-1396.30 

1519.99-1522.88 



5.2-10.8/S 
2.9-9.4/L 



46 



Calculated mineral content of selected samples 
from potassium-bearing Intervals with summation 
of percent K.O as ore mineral 



Drillhole No.: Drillhole designations; P, Energy Research and Development Adminis 
FC, Farm Chemical Res. Dev . Corp; IMC, International Minerals and Chemical Corp; 
NFU, Farmers Edu, and Coop. Union of America; D. Duval Sulphur and Potash Co.; 
U, U.S. Potash Co, , Inc. 



Calculated Minerals Present and other headin 
C. carnallite; Gl , glaserite; Gu, glauberit 
Le, leonite; Lo , loeweite; S, sylvite; Va , 



chofite; Bl, bloedite; 
Ite; L, langbeinite; 



Drillhole Ore Sample Depths of 
No. Zone No. interval 
(feet) 



;d minerals present 
5ight percent) 



Polyhalite Halite Sylvite Langbeinit 



K.O as Weighted average 
ore K as ore mineral 
Other minerals for intervals preceding 
ninerals (percent) by ore zone 

(feet and percent) 



-104 3- 
3 



1527.50-1528.90 1.40 

1528.90-1530.50 1.60 

1539.50-1540.20 1.20 

1319.58-1321.25 1.69 

1321.25-1322.83 1.58 

1361.10-1362.18 1.08 

1362.17-1364.50 2.33 

1364.50-1366.50 2.00 

1366.50-1367.42 0.92 



5 
5 
5 


1406.75-1409.42 
1409.42-1410.00 
1410.00-1411.66 


2.66 
0.58 
1.66 


4 


1471.66-1474.00 


2.33 


3 

3 
3 


1484.91-1487.33 
1487.33-1490.30 
1490.25-1491.33 


2.42 
2.92 



2.0/Ka 

0.7/C 3.79/S 
12.5/Ka 17.22/S 



14.44/L 
8.03/L 
7.05/L 



2.9-11.4/S 
mixed ore equivalent: 
2.9-6.66/L 



3.27-10.28/S 



3.6/L 
1.86/L 
8.7/L 



3.41-11.6/S^ 
(4.0-10.0/S)- 



4.9-11.2/L 
2.33-8. 54/L 



Data provided by U.S. Geological Survey. 



1/ Trace amount; equals to 2.0% 

II Incomplete dissolution of sample 

3/ 5.97. Insolubles, by weight 

4/ Incomplete or unreliable assay 

5/ Grade adjusted to 4 foot interva 

6/ High insoluble content 



7/ 


7.1% Potassium Assa 


8/ 


Outside of the ERDA 


9/ 


Raw data unavailabl 


0/ 


Company interval da 



feet , included du 
company figures 



47 



R30E 



R^^ 




DATA PROVIDED BY USGS CONSERVATION QIMC-448 
DIVISION ANDERDA 

LEGEND 

© Potash drill holes 

a ERDA potash drill holes 

p^ Measured and indicated mineralization 

I I Federal surface and mineral rights 

l,',l'i State surface and mineral rights 

F.'^J Private surface and mineral rights 

F^ Privote surface, all mineral rights owned 
■^•^ by Federal Government 

c^ Private surface and mineral rights, except 
^^■^ oil and gas federally owned 

^— Proposed WIPP site outline 

Zone boundaries and areas provided by ERDA 



Measured and indicated mineralization are at a cut off 
of 8 % K2O as langbeinite or 14.0% KgO as sylvite 
or equivalent grade of mixed langbeinite-sylvite 
occurring in a minimum 4 foot interval 

Zone 
I - 58 acres 
n -1,889 acres 
in - 6,201 acres 
m - 10,812 acres 



FIGURE 18. — Composite map of mineralization in various ore zones at 8 and 14 percent KgO as 
langbeinite and sylvite, respectively. 



48 



R^^ 




DATA PROVIDED BY USGS CONSERVATION &IMC-448 
DIVISION AND ERDA 

LEGEND 

@ Potash drill holes 

la ERDA potash drill holes 

Kyi Measured and indicated mineralization 

I I Federal surface and mineral rights 

t,,','i State surfoce and minerol rights 

r,"!^ Private surface and mineral rights 

P^ Private surface, all mineral rights owned 
^^^ by Federal Government 

FT^ Private surfoce and mineral rights, except 
^ oil and gas federal ly owned 

^—Proposed WIPP site outline 

Zone boundaries and areas provided by ERDA 



Measured and indicated mineralization are at o cut off 
of 4 % K2O OS langbeinite or 10.0% KgO as sylvite 
or equivalent grade of mixed langbeinite-sylvite 
occurring in a minimum 4 foot interval 

Zone 
I - 58 acres 
n -1,889 acres 
m - 6,201 acres 
nr - 10,812 acres 



FIGURE 19. — Composite map of mineralization in various ore zones at 4 and 10 percent K2O as 
langbeinite and sylvite, respectively. 



49 



R30E 



m^ 




DATA PROVIDED BY USGS CONSERVATION OIMC-44B 
DIVISION AND ERDA 

LEGEND 

© Potosh drill holes 

B ERDA potash drill holes 

p^ Measured and indicated mineralization 

I I Federal surface and mineral rights 

m State surface and mineral rights 

r,.,l Private surface and mineral rights 

p-i Private surface, oil mineral rights owned 
^^^ by Federal Government 

p?^ Private surface and mineral rights, except 

^''^ oil and gas federally owned 
^— Proposed WIPP site outline 
Zone boundaries and oreos provided by ERDA 



Measured and indicated mineralization are at a cut off 
of 3% KgO OS langbeinite or 8.0% K2O as sylvite 
or equivalent grade of mixed langbeinite-sylvite 
occurring in a minimum 4 foot interval 



Zone 
I - 58 acres 
n -1,889 acres 
nr - 6,201 acres 
nr - 10,812 acres 



FIGURE 20. — Composite map of mineralization in various ore zones at 3 and 8 percent KgO as 
langbeinite and sylvite, respectively. 



50 



FINANCIAL ANALYSIS 



METHODS USED 

Present value analyses were used to determine 
the worth of the potash mineralization in the 
WIPP site. These are the values that would be 
chargeable to the WIPP facility if it were built 
and the potash products were not recovered. All 
reserves (presently commercial) and presently 
subeconomic resources of potassium mineral 
deposits were examined. Incomes, costs, and in- 
vestments are assumed to occur as discrete 
amounts over the life of the project. For eco- 
nomic evaluation purposes and to compare 
equivalent values that have been adjusted for 
the time value of money, these amounts are con- 
verted (discounted at an appropriate discount 
rate) to equivalent values (1977 dollars) at one 
point in time (project initiation date). Present 
value of a mineral deposit is defined for these 
analyses as the present worth of the cash flows 
from a hypothetical potash operation minus the 
present worth of the investments (23). The po- 
tash values are taxes, royalties, and bonus bid 
amounts that would be generated from potash 
product sales and paid to the several levels of 
government. Such taxes, royalties, and bonus 
bid amounts (or their equivalent value) will not 
accrue to the governments if the potash deposits 
are not developed. All other values of produc- 
tion and investments originate when projects are 
initiated, or when capital is invested, and there- 
fore are not a loss if the unmined, commercial 
potash is not developed. 

Cash Flow Estimates 

The cash flows are the gross revenues minus 
all direct, indirect, overhead, and front office 
production costs. These do not include book- 
keeping charges for depreciation, depletion, 
amortization and extraordinary charges to cap- 
ital reserves that are not actually paid out. 

Cash flows of the project are the gross reve- 
nues minus (1) mine operating costs; (2) mill 
operating costs; (3) all overhead costs, including 
front office charges; (4) all Federal and State 
royalties; and (5) all Federal and State taxes. 

Cash flows are the project's total self-gener- 
ated cash after costs and are the amounts that 



provide for the return of the original investment 
plus the interest on that investment. The cash 
flows are a production-cost element, in dollar 
amounts generated by the project, that return 
the investment plus sufficient interest on the 
investment to induce investors to provide the 
capital for the project. 

Investment Estimates 

The total investments include (1) exploration 
and engineering study; (2) acquisition costs (in- 
cluding any lease bonus payment); (3) mine 
preparation; (4) mine plant investment; (5) mine 
equipment; (6) mine reinvestments in capital 
items during project life; (7) mill plant invest- 
ment; (8) mill capital reinvestments during proj- 
ect life; (9) all working capital; and (10) startup 
and break-in costs. 

The cash flows and investments are dis- 
counted by continuous discounting (versus dis- 
crete discounting) at an interest rate ( 1 5 percent) 
suitable to investors in the potash industry. A 
lower interest rate (return on investment) is as- 
sumed to be unacceptable and would divert the 
capital to more attractive investment opportun- 
ities. 

The values in the commercial potash deposits 
that would be foregone in favor of the instal- 
lation of the Waste Isolation Pilot Plant include 
the amounts of bonus bids that could be paid by 
potash investors to the Federal and State gov- 
ernments for potash leases. In addition to these 
amounts are the present values of Federal and 
State royalty payments and taxes that would 
have been generated in the production of the 
potash products. To estimate the present value 
of the taxes and royalty payments, these amounts 
were discounted at 8 percent interest from the 
year in which they would have occurred to 1977 
dollars. The total values that would be foregone 
are (1) bonus bid amounts, (2) present value of 
tax amounts, and (3) present value of royalties. 

In the event the WIPP installation was built, 
and if the government-owned potash in the site 
were under lease, these values would be fore- 
gone to the government entities. In addition, 
the lessee would lose any amounts, including 
interest, that have been paid out for acquisition 



51 



of the property, for exploration, for develop- 
ment engineering studies, for legal fees, etc. If 
presently commercial potash deposits are leased, 
the lessee will also lose a potash investment op- 
portunity in addition to the aforementioned out- 
of-pocket amounts spent. 

Determination of Commercial and 
Subeconomic Mineralization 

When the sum of cash flov/s, discounted at 15 
percent interest, equals or exceeds the sum of 
the investment discounted at 15 percent interest, 
both to 1977 dollars, the project is considered 
to be commercial; and the potash deposit is class- 
ified as a reserve or ore. When the sum of the 
discounted cash flows is less than the sum of the 
discounted investments, the project is not com- 
mercial, and the potash deposit is considered to 
be economically submarginal and is classified as 
a resource. These potash resources may become 
ore at some future time when potash becomes 
more valuable relative to costs of production. 

When the cash flows will return the original 
investment in a mine-mill complex and more 
than the acceptable 15-percent interest rate, the 
potash deposits being examined have a present 
value or worth in excess of all other investments. 
In this case, the owners of the potash (in the 
WIPP site, the Federal and State governments) 
could require payment of the present value or 
worth in the form of a bonus bid amount for a 
lease to produce the potash (designated an ac- 
quisition cost in this report). 

To examine the potash resources (presently 
subeconomic deposits), the market prices of the 
products were increased, without increasing the 
costs of production, until the deposits became 
commercial. This analysis provides a guide to 
the potential value of the potash resources 
within the WIPP site. An estimate of the time 
at which these increased potash values might 
occur is so speculative it is not made in these 
analyses. 

An analysis of the impact of the loss of the 
potash reserves and resources, such as the effect 
on the gross national product and balance of 
payments, was not made because it was beyond 
the scope of this study. 

Taxes and Royalties 

Taxes are a major cost item, and a discussion 
of applicable New Mexico State taxes is in- 
cluded. New Mexico potash operations are sub- 
ject to the following taxes: Federal corporation 
income tax; New Mexico corporation income 
tax; resource, processors, and service tax; sev- 



erence tax; property tax; and in lesser amounts — 
franchise tax, corporate organization and qual- 
ification fees; motor vehicle registration fees; 
motor carrier fees; and other indirect taxes. 

The State Corporation Commission adminis- 
ters and collects the corporation annual report 
filing fees, the corporation and qualification 
fees, the franchise tax, miscellaneous motor car- 
rier fees, and the pipeline companies tax. The 
Bureau of Revenue administers the resource 
excise tax, the severance tax, and the gross re- 
ceipts tax; the Property Tax Department ad- 
ministers and collects the mining property tax. 

Federal Corporation Income Tax 

Federal income tax for corporations is based 
on gross income minus certain deductions. 
These deductions in general are costs of oper- 
ations and include compensation of officers, sa- 
laries and wages, repairs, bad debts, rents, taxes, 
interest, contributions, amortization, deprecia- 
tion, depletion, advertising, pension, profit- 
sharing, and employee benefit programs. Fed- 
eral income tax rates for corporations are 20 
percent on the first $25,000 of taxable income, 
22 percent on the next $25,000, and 48 percent 
on all taxable income over $50,000. 

New Mexico Corporation Income Tax 

The New Mexico income tax is based on the 
Federal corporation income tax. The State in- 
come tax is 5 percent of the net taxable income 
within the State of New Mexico. 

If the corporation operates exclusively in New 
Mexico, interest on U.S. obligations, to the ex- 
tent that they are included in Federal taxable 
income, and interest on amounts taxed as in- 
come to another member of an affiliated group 
of corporations are listed as nontaxable income 
and are deductible. 

If the corporation operates both within and 
outside the State, the taxable income for the 
State of New Mexico is estimated by dividing the 
nonbusiness and business income between New 
Mexico and elsewhere. 

Nonbusiness income includes dividends, in- 
terest, rents, royalties, profit or loss on the sale 
of nonbusiness assets, and partnership income. 
The nonbusiness income is allocated to New 
Mexico in three ways: (1) to the extent that it is 
generated in the State; (2) if the nonbusiness 
income was earned from the sale of property, 
if the property is located in New Mexico at the 
time of sale; and (3) if New Mexico is considered 
the commercial domicile of the corporation. 

The division of all business income is appor- 



52 



tioned to New Mexico by a three-factor formula: 
(1) the property factor is the average value of 
real and tangible personal property owned or 
rented and used in New Mexico divided by the 
total real and tangible property owned, rented, 
or used by the corporation. The average value 
of the property owned is determined by aver- 
aging property values at the beginning and the 
end of the year; the value of property rented is 
eight times the net annual rental rate; (2) the 
payroll factor is the amount paid as compensa- 
tion by a corporation to its employees in New 
Mexico divided by the total compensation paid 
by the corporation in the tax period. Compen- 
sation means wages, salaries, commissions, and 
other forms of remuneration for personal serv- 
ices except wages paid to independent contrac- 
tors; and (3) the sales factor is determined by 
dividing all gross receipts of the corporation 
from transactions and activity in the regular 
course of business in New Mexico during the 
tax period by the total sales of the corporation 
in the same period. Sales in this context exclude 
nonbusiness income. The three factors are then 
averaged to determine the percentage of the 
corporation's business income that is applicable 
to New Mexico taxes. 

Finally, that portion of the nonbusiness in- 
come allocated to New Mexico is added to the 
apportioned business income to determine the 
net taxable income for a corporation operating 
in the State. 

Resources, Processors, and Service Tax 

Resources, processors, and service tax are 
mutually exclusive, that is, only one of the taxes 
is imposed on the potash produced. 

The resource tax applies to unprocessed 
products sold. The amount taxable is deter- 
mined by subtracting from the gross value of 
the products sold (1) gross sales to any govern- 
ment entity, (2) government royalty payments, 
and (3) any service charge on which a service 
tax is payable. This taxable value is then mul- 
tiplied by one-half of 1 percent to arrive at the 
resource tax payment. 

The processors tax is based on all processed 
products sold. The taxable value is determined 
by subtracting from the gross value of the proc- 
essed products sold, the same deductions ap- 
plicable to the resource tax. This value is then 
multiplied by one-eighth of 1 percent to arrive 
at the processors tax payment. 

The service tax is imposed on a corporation 
that either mines or processes New Mexico po- 
tash owned by another company or not other- 
wise taxed with a resource or processors tax. 



The rate applied is the same as the rate used to 
calculate the resource or processors tax. 

Severance Tax 

The New Mexico severance tax is an excise 
tax based on an adjusted gross value of potash 
at the first marketable point. 

The adjusted gross value is 33-'/* percent of 
the proceeds realized from the sale of muriate 
of potash and sulfate of potash magnesia, in 
terms of standard grades, and 33- '/a percent of 
products consumed in production of other pot- 
ash products. The taxable value is equal to the 
adjusted value minus all royalties and up to 50 
percent of expenses incurred. 

In the case of mine-run salt that has a posted 
market price at the point of production, the tax- 
able value is 40 percent of gross value minus all 
royalties and expenses incurred in hoisting, 
crushing, and loading. 

The tax rate is 2- 'A percent of the taxable 
value. 

Property Tax 

Property tax is imposed on all property used 
in connection with the production of potash. 
The tax is based on the property value of im- 
provements and production. The property value 
of improvements, for tax purposes, is deter- 
mined to be equal to the gross value of the prod- 
ucts sold minus any royalties paid. The property 
value of the orebody, for tax purposes, is de- 
termined to be one-half the gross value of the 
products sold minus any royalties paid. The sum 
of the property value of improvements and pro- 
duction is then multiplied by a 33-'/3 percent 
assessment ratio to determine the taxable value. 
The taxable value is multiplied by the mill levy 
of the district in which the property is located 
to determine the property tax due the county. 
The mill levy assessment for the Carlsbad School 
District, Eddy County, was 19.8 mills in 1976. 

Rents and Royalties on Federal Leases 

Royalties on Federal land leases are negoti- 
ated on an individual case basis prior to issuance 
of the lease. Royalty rates range from 2 to 7- 
'/2 percent of the gross value of production. 
Current lease agreements require a payment of 
royalty on a minimum annual production be- 
ginning the sixth full calendar lease year. 

Rental for potassium leases is 25 cents per acre 
(62 cents per hectare) for the first calendar year; 
50 cents for the second, third, fourth, and fifth 



53 



year; 1 dollar for the sixth and succeeding years. 
The rent is credited against the first royalties 
paid during the year for which rent was paid. 
In any one State, a lease may not exceed 2,560 
acres (1,036 hectares). Also, holdings in leases 
may not exceed 25,600 acres (10,360 hectares) 
in one or more mining units. 

If conditions warrant, the Secretary of the In- 
terior can reduce the rental or minimum royalty. 



Rents and Royalties on State Leases 

Royalties on State lands are payable on a ne- 
gotiated basis but not less than 5 percent of the 
sale value of the minerals produced then deliv- 
ered to the nearest shipping point. Rental 
charge is negotiated but not less than 10 cents 
per acre (25 cents per hectare), with a minimum 
first year rental of $100. 



ESTIMATION OF CURRENT MINE AND 
MILL CAPITAL AND OPERATING COSTS 

In the financial evaluation of mineralized 
langbeinite and sylvite zones found within the 
WIPP site boundaries, the costs for three types 
of currently operating potash mining properties 
were estimated. The capital and operating costs 
estimated were then extrapolated to estimate 
costs of mines and processing plants of a size 
and design best suited to recover the deposits 
determined to be in the WIPP site. Personnel of 
the System Operations Group (SOG) and the 
Domestic Evaluation Group of the Minerals 
Availability System (MAS) toured seven bene- 
ficiation plants and underground potash oper- 
ations in the Carlsbad area to evaluate potash 
mine and mill operations. 

From the information gathered, operating 
and capital replacement costs of current potash 
mining and milling properties were estimated 
for operations applicable to WIPP site ores. Con- 
sideration was given to the economics of (1) us- 
ing existing mill complexes with a new mine 
operation and (2) building a new mine-mill com- 
plex. 



Operations for WIPP Study 

Cost estimates were developed for a 13,000- 
ton-per-day (11,794-metric-ton-per-day) oper- 
ation, a 7,000-ton-per-day (6,350-metric-ton-per- 
day) operation, and an 8,500-ton-per-day (7,- 
711-metric-ton-per-day) operation. A brief dis- 
cussion of these three operations follows. 



13,000-ton-per-day Mine-Mill 

Estimates were made of a mining operation 
with a design capacity of about 13,000 tons per 
day (11,794 metric tons per day) to extract 
mixed ore (sylvite-langbeinite) and langbeinite 
ore (fig. 21) using a room-and-pillar mining sys- 
tem (figs. 22 and 23). Conventional mining (figs. 
24 and 25) is used because of hardness of lang- 
beinite and polyhalite minerals. 

Entry is by vertical shafts with connecting 
drifts. Mobile trackless equipment is used to 
drill, load, and haul the ore to a conveyor system 
near the working areas. The blasting agent used 
is ANFO. Ore is transferred by conveyor to the 
production shaft and hoisted to the surface. 

This operation includes a beneficiation system 
that recovers sylvite and langbeinite from 10,000 
tons per day (9,072 metric tons per day) of 
mixed ore by heavy media separation and sylvite 
flotation. The mill is designed for ore having a 
clay content of 1 to 1.5 percent. In addition, the 
operation includes recovery of langbeinite from 
a 3,000-ton-ore-per-day (2,722-metric-ton-ore- 
per-day) langbeinite wash circuit. Sylvite and 
langbeinite fines are then used as feed material 
in the production of potassium sulfate (fig. 17). 
The sulfate plant was designed for an output 
capacity of 500 tons per day (454 metric tons 
per day) of product (fig. 26). 



7,000-ton-per-day Mine-Mill 

The 7,000-ton-per-day (6,350-metric-ton-per- 
day) room-and-pillar system includes the mining 
of two separate mineralized zones, one sylvite 
and the other langbeinite (fig. 27). The sylvite 
mining operation has a capacity of 4,000 tons 
per day (3,629 metric tons per day), while the 
langbeinite operation has a 3,000-ton-per-day 
(6,350-metric-ton-per-day) capacity. 

Vertical shafts and connecting drifts provide 
access to each mineralized zone. Both mines 
utilize conventional room-and-pillar mining 
methods; drilling, blasting, and loading are 
done by mobile trackless equipment. A con- 
veyor-belt system transports the ore to the pro- 
duction shaft. The ore is then hoisted to the 
surface and transported to the mill. Sylvite is 
recovered from the low-clay (1 to 1.5 percent) 
ore by flotation, and langbeinite is beneficiated 
from its ore by water leaching of the impurities. 
Sylvite and langbeinite fines are then used as 
feed material in the production of potassium 
sulfate (fig. 17). This sulfate plant has an output 
design of 225 tons per day (204 metric tons per 
day) of product. 



54 




Crushing 
& sizing 



100 200 400 

liinl I I 



FIGURE 21.— Generalized layout of 13,000-tpd surface plant. 



55 



o o o 

.-. "O T- 



^11 



1 n rnrnr 



3 



PS 



tinnnnn 

• > ^ > > > > 



ij 



56 




pn/D 



nnn n 



D 



DO-QO-n 




57 




2-boom jumbo, 1 man, 10-foot hole depth, 
1-3/4-inch-diam hole, 45 min/round 



DRILLING 



KKKKKKKK^^^^ 



VSSVsSSSSSSSSVsSSS'C 



UNDERCUTTING 




1 man, 10-foot length, 6-inch thick across 
floor of panel, 45 min/round 



1 man, 16 primers, 16 caps, and 65 lb ANFO, 
30 min/round 



LOADING AND BLASTING 




1 loader operator 

2 shuttle-car operators 
45 min/round 



LOADING 



fa 



1 operator 

5 roof bolts/face 



//////////////// ///y ////// ////////////y 

ROOF BOLTING 



Approximate dimension of round - 5'xl0'x26', 
thickness of bed varies. 

Tonnage - 88.6 tons/round, varies 
as thickness of bed. 

FIGURE 24. — Diagram of conventional mining sequence. 



8,500-ton-per-day Mine-Mill 

The third operation is an 8,500-ton-per-day 
(7,700-metric-ton-per-day) room-and-pillar mine 
(fig. 27) using, in part, a continuous mining sys- 
tem (fig. 28). The processing system is amenable 
to sylvite ore with a low langbeinite and poly- 
halite content. Approximately 75 percent of the 
mine output is extracted by continuous mining 
methods. In the continuous mining method, 
boring machines are used to extract and load 



the ore into shuttle cars which transport the 
mined ore to a conveyor system. The conveyor 
system transfers the mined ore to the shaft area 
to be hoisted to the surface. 

The remaining 25 percent of the ore is mined 
by conventional mining method (fig. 24) much 
in the same manner as described previously. 

The beneficiation system will recover sylvite 
from the ore by solution-crystallization methods 
(fig. 15). The ore is assumed to contain approx- 
imately 4 percent clays. 



58 




FIGURE 25. — Typical round dimensions for conventional mining. Approximate dimensions of 
broken round — 5 by 26 by 9.5 ft (hole length 10 ft, breaking length 9.5 ft); volume, about 1,235 
cu ft; 16 holes per round; hole diameter, iy4 in; about 7 ft of ANFO per hole; 2 delays — and 1 
millisec; tonnage 90 tons/round, assuming a powder factor of 0.7 lb explosive per ton of ore; 0.82 
lb of ANFO per foot of drill hole = 90 lb of ANFO per round; primer — 1- by 8-in 70-percent 

gelatin. 



Mine Capital Costs 

Mine capital investments are those costs as- 
sociated with bringing an operation into pro- 
duction. The estimate is based on 1977 dollar 
values and is the cost for new (replacement) 
mine-mill complexes. Mine capital investments 
include exploration, development, mine plant, 
mine equipment, and working capital. 

Exploration investments include the cost to 
define the mineralized zone in size, tonnage, 
and grade. This is done by estimating the cost 
of the drilling program and of a geologic and 
engineering analysis of the data obtained in the 
study area. 

The development costs include all invest- 
ments involved in preparing the operation for 
production. This includes the cost of construct- 
ing roads and railroads and providing utilities 
to the study area. The development costs also 
include all investments related to sinking of 
shafts, development of main haulage drifts, and 
preparing the working areas for full production. 

Mine plant costs include all investments as- 
sociated with permanent structures such as 
buildings, shops, and permanent systems such 
as ventilation and hoisting. These costs are con- 
sidered to last the life of the operation. 



Mine equipment costs include all costs asso- 
ciated with mobile underground equipment as 
well as support equipment on the surface. Other 
items such as tools, personnel-carrying vehicles, 
and an inventory of mining supplies are also 
included in this investment. 

Working capital is the amount of money 
needed to account for all operating expenses 
between the time of first production and the 
time of initial revenue from the products. This 
cost was estimated to be equal to 6 months' op- 
erating cost for each mining system. 

A summary of the estimated capital invest- 
ments of the three potash operations is given in 
table 14. 



Mine Operating Costs 

Mine operating costs are the expenses in- 
curred in extracting the ore and waste rock from 
the deposit. Operating costs are subdivided into 
three basic subcategories: (1) direct, (2) indirect, 
and (3) fixed costs. Factors that influence op- 
erating costs include hardness of rock, mining 
thickness, and variability of mineralization. 

Direct costs are the investments related to the 
actual mining output. These costs include direct 





TABLE 14 


. — Summary of estimated mine capital investments 


Mine 


Plant cost 


Equipment 


Exploration 

and 
development 

cost 


Working 
capital 


Total 
investment 


13,000 tpd 


$6,093,500 
9,015,200 
5,712,500 


$ 8,114,400 
9,817,900 
11,691,300 


$13,248,300 
15,079,400 
10,278,100 


$3,906,200 
4,330,600 
4,659,800 


$32,362,400 


7,000 tpd 

8,500 tpd 


38,243,100 
32,341,700 



59 




Production 

shaft and 

mine head frame 



Service shaft 
and mine D 
headframe 



Scale, feet 



FIGURE 26. — Generalized layout of 7,000-tpd surface plant. 



60 




Change room 




A 


1 1 


Mine and ^^^ 




ill hoist \/ y \ 


./ 


Hoist house 


X 


^o' 


■^Ore hois 



Scale, feet 

FIGURE 27.— Generalized layout of 8,500-tpd surface plant. 



mine labor and supervision, maintenance labor 
and supervision, maintenance supplies, and 
material and utilities costs. 

Indirect costs are related to support functions 
such as administrative technical labor, mainte- 
nance of the surface facilities, and general over- 
head. The direct and indirect costs would be 
affected by fluctuations in mine output, varia- 



bility in rock and mineralized composition, and 
external economic factors. 

Local taxes and insurance make up the fixed 
costs. These costs remain relatively constant and 
are not affected by physical or economic changes 
to the mining system. 

A summary of the estimated operating costs 
for the three typical potash operations is shown 
in table 15. 



61 



O 2 



i 



s- 

S- 
CQ 

I 



^ ° C 

sl-s 

c/5 3 £ 

2 u u 

3 S dj 

S c 3 

.£ ■?i oJ "a; 
S o.S-^ 
55 ^ u o 

i 2 bog 

(J o-§ C 

bC°- « 
C |2 "3 bC 

=2 id -^ j= 

1^ -if 3 

^ ="ii >^ 

bo2 =" 5 
'S jj c o 

w^ >- ii o 

G OS O ^ 

°2 

(M a, ° 
. , o <u 



T 2 ■- 
I 2 u 



62 



TABLE 15. — Summary of estimated mine 
operating costs 

(Dollars per short ton of materials mined) 



Mine 


Direct 


Indirect 


Fixed 


Total 


13.000 tpd 

7,000 tpd 

8,500 tpd 


1.80 
3.00 
2.97 


0.31 
.79 
.35 


0.05 
.10 
.10 


2.16 
3.89 
3.42 



Mill Capital Costs 

The mill capital investments for the three op- 
erations include the cost for the plant, equip- 
ment, and working capital. The plant and 
equipment investment includes all costs associ- 
ated with the permanent structures such as 
buildings and labs, processing equipment, and 
support systems such as piping, insulation, etc. 
The working capital is the amount of money 
needed to account for all operating expenses 
between the time of first production and the 
time of initial revenue from the products. These 
costs were estimated to be equal to 6 months' 
operating cost for each mill system. 

Mixed Ore Plant 

The mixed ore plant has a design capacity of 
about 10,000 tons per day (9,072 metric tons per 
day) of a mixed (sylvite-langbeinite) low clay (1 
to 1.5 percent) ore. This ore is separated into 
sylvite and langbeinite products and fines are 
fed into a sulfate plant designed to produce 500 
tons per day (464 metric tons per day) of prod- 
uct. Langbeinite ore is processed in a 3,000-ton- 
per-day (2,722-metric-ton-per-day) langbeinite 
leach circuit. 

The three mill sections of this plant were eval- 
uated separately so that any portion of the plant 
could be used where variations from the models 
were needed for WIPP ore. In addition, the 
1 0,000-ton-per-day (9,072-metric-ton-per-day) 
mixed-ore circuit was costed in two parts: a syl- 
vite flotation circuit and a langbeinite float- 
heavy media circuit, allowing each circuit to be 
used separately if the WIPP site ore was a lang- 
beinite or sylvite ore rather than a mixed ore. 

Sylvite Ore Flotation Plant 

The sylvite flotation plant has a design ca- 
pacity of about 4,000 tons per day (3,629 metric 
tons per day) of a low-clay (1 to 1.5 percent) 
sylvite ore beneficiated by flotation. The con- 
centrate fines from this operation and the fines 
from a 3,000-ton-per-day (2,722-metric-ton-per- 
day) langbeinite leach circuit concentrate are fed 
to a sulfate plant designed to produce 225 tons 
per day (204 metric tons per day). As with the 



TABLE 16. — Summary of estimated plant 
capital investments 

Plant description Costs 
13,000-tpd mill complex: 

10,000-tpd mixed (sylvite-langbeinite) ore circuit trmploying 

heavy media separation flotation _ $26,789,400 

3,000-tpd langbeinite leach circuit 4,748,100 

Sulfate plant— 500 tpd product _ 24,161,700 

Total __ _ $55,708,200 

7,000-tpd mill complex: 

4,000-tpd sylvite flotation circuit $20,325,500 

3,000-tpd langbeinite leach circuit 5,407,200 

Sulfate plant— 225 tpd product ___ _ 15,242,800 

Total __ ___ _ $40,975,500 

8,500-tpd mill complex: 

8,500-tpd sylvite crysuUization plant $52,899,500 



mixed ore plant, the three mill sections of this 
plant were evaluated separately so that any por- 
tion of the plant could be used if variations from 
the models were needed to evaluate a hypo- 
thetical plant for WIPP deposits. 

Sylvite Ore Crystallization Plant 

A sylvite crystallization plant was designed to 
beneficiate 8,500 tons per day (7,7 1 1 metric tons 
per day) of a sylvite ore containing approxi- 
mately 4 percent clays by dissolution and re- 
crystallization of the sylvite. 

The three mills have separate mill circuits for 
different ore treatment in the 10,000- and 7,- 
000-ton-per-day operations and for potassium 
sulfate production. Each circuit was evaluated 
separately so that any portion of the plant could 
be used for estimating capital and operating 
costs of mine-mill complexes best suited for re- 
covery of the WIPP deposits. 

A summary of the estimated capital invest- 
ments is shown in table 16. 

Mill Operating Costs 

Mill operating costs are the expenses incurred 
in recovering the potash from the mined ore. 
Operating costs are subdivided into three basic 
subcategories: (1) direct, (2) indirect, and (3) 
fixed costs. Direct costs are those related to the 
actual production such as labor, water, electric- 
ity, and reagent requirements. Indirect costs are 
expenses related to support functions such as 
administration, technical labor, maintenance of 
the facilities, and general overhead. The direct 
and indirect costs are affected by mill input, var- 
iability of the rock, mineral composition, and 
external economic factors. Local taxes and in- 
surance comprise the fixed costs; these costs re- 
main relatively constant. A summary of the 
estimated operating costs for the three mill com- 
plexes is shown in table 17. A separate cost is 
shown for the different ore treatment circuits 
within the mill complex. 



63 



TABLE 17. — Summary of estimated annual 
mill operating costs 

Plant desmplton Aimual rml\ 

I3.000-tpd mill complex: 

10,000-tpd mixed (sylvite-langbeinile) ore circuit employing 

heavy-media separation flotation $9,755,900 

3.000-tpd langbeinite leach circuit 1.550,400 

Sulfate plant— 500 tpd product 5,563,000 

Total $16,869,300 

7,000-tpd mill complex: 

4,000-tpd sylvite flotation circuit $5,038,600 

3,000-tpd langbeinite leach circuit 1.498,200 

Sulfate plant— 225 tpd product 2,267,500 

Total ____ $8,804,300 

8.500-tpd mill complex; 

8.500-tpd sylvite crystallization plant ___ $8,294,000 



Utility (Infrastructure) Costs 

Utility costs include estimated costs of re- 
quired roads, railroad spurs, water supply sys- 
tems, electric utility lines, and natural gas supply 
lines. 

Water 

The water supply for a new potash-producing 
mine-mill complex is a critical item. Possible 
sources of water considered for potash refining 
are the Pecos River and aquifers in the Capitan 
Reef, Rustler, and Ogallala Formations. The 
best quality water found in the region is pro- 
duced from the Caprock area of the Ogallala 
Formation, 25 to 30 miles (40 to 48 kilometers) 
to the northeast of the study area. This source 
presently supplies the potash industry with 
about two-thirds of its water and is being used 
increasingly by the industry. The water quality 
is good, ranging from 500 to 600 parts per mil- 
lion total dissolved solids. 

This study assumes that Caprock area-Ogal- 
lala Formation water will be available for a new 
refinery. Adequate water rights are available 
and well site leases are obtainable on most of the 
Caprock field. The State of New Mexico owns 
most of the producing area. Present wells are 
200 to 250 feet (61 to 76 meters) deep and are 
spaced on about '/4-mile (0.4-kilometer) centers. 
Each well, if properly designed and developed, 
can produce 200 to 250 gallons per minute (756 
to 945 liters per minute) for 30 to 40 years. It 
is estimated that a pipeline from the Caprock 
field to a new refinery in the study area would 
be 30 miles (48 kilometers) long. A 600-gallon- 
per-minute (2,268-liter-per-minute) water sys- 
tem is estimated to cost about $3.5 million in- 
stalled, a 3,600-gallon-per-minute (13,608-liter- 
per-minute) water system about $6.9 million. 



Electric Power 

Electric power is supplied in the area by the 
Southwestern Public Service Co. A typical load 
for a potash mine and refinery is estimated to 
be about 6,000 kilowatts, with usage at approx- 
imately 4 million kilowatt-hours per month. All 
power supplied to the potash industry is by 69- 
kilovolt powerlines. These lines are estimated to 
cost approximately $35,000 per mile. The volt- 
age is reduced at a substation near the mine or 
refinery. The present rate for power is esti- 
mated to be approximately 2 cents per kilowatt- 
hour. A 69-kilovolt line to the Kermac opera- 
tion, capable of handling the added capacity re- 
quired for a new mine and mill complex, could 
be extended into the study area. 

Natural gas for a refinery could be supplied 
by the Gas Company of New Mexico. A 6-inch 
(15.2-centimeter) line presently supplies the 
Kermac refinery and has sufficient capacity to 
supply an additional refinery. The normal gas 
usage by existing refineries ranges from 15,000 
to 100,000 million Btu per month. The present 
intrastate commercial rate for gas in New Mex- 
ico is estimated to range from $ 1 . 1 5 to $ 1 .35 per 
million Btu. To supply a new refinery, a 4-inch 
(10.2-centimeter) pipeline from the Kermac op- 
eration south to the study area would cost ap- 
proximately $27,000 per mile. 

All commercial natural gas customers except 
Kermac have alternate emergency supplies of 
burner fuel to use in case of gas curtailment. 
Kermac stores propane for an alternate emer- 
gency fuel. Burner fuels are stored in 20,000- 
to 30,000-gallon (75,600- to 113,400-liter) stor- 
age tanks at the potash refineries. Estimated 
costs for alternate fuels are $2.36 per million 
Btu for No. 2 burner fuel and $2.72 per million 
Btu for propane. 

Coal could be used instead of natural gas in 
the refining of potash. Steam could be produced 
in coal-fired boilers and its heat used for the 
leaching, drying, and granulating steps of the 
refining process. It is estimated that the extra 
capital cost to build a coal-fired refinery over 
the cost of building a natural-gas-fired refinery 
is 15 to 25 percent. Coal for use in the Carlsbad 
area probably could be supplied from a coal 
mine near Gallup, N. Mex. It is estimated that 
this 12,000 Btu per pound (26,400 Btu per kil- 
ogram) of coal would cost about $0.78 per mil- 
lion Btu delivered. Table 18 shows a comparison 
of fuel costs between coal and natural gas for a 
langbeinite wash plant. 

Table 19 shows a comparison of capital in- 
vestment necessary to accommodate natural gas 
and coal as fuel for a langbeinite wash plant. 



64 



TABLE 18. — Comparison of fuel cost for a 
3,000-tpd langbeinite wash plant 



Fuel 


Cosl/MM Blu 


Total 
MM Btu/year 


Total 

cost/year 


Natural gas 

Coal .. 


$1.15 

.78 


160,313 
160,313 


$184,360 
125,040 



TABLE 19. — Comparison of capital 
investment: coal versus natural gas in 
3,000-tpd langbeinite wash plant 



Fuel utilized 


Percent 
markup 


Base price 


Total cost 


Additional 





15 
20 
25 


$4,748,100 
4,748,100 
4,748,100 
4,748,100 


$4,748,100 
5,460,300 
5,697,700 
5,935,100 




Coal .. 

Coal 

Coal 


712,000 
949,600 
1,187,000 



The savings that would accrue by using coal 
instead of natural gas is $59,320 per year. The 
dollars of fuel saving, as simple interest on the 
additional investment necessary to use coal, pro- 
vide 8.3 percent interest on the additional in- 
vestment for the best use and 5 percent for the 
most costly use. 



Access to the Area 

Access into the area would be north from State 
Highway 128 or south from U.S. Highway 180. 
If an all-weather road were to be constructed 
into the WIPP site, it probably would be built 
south from Highway 180, which is better suited 
to heavy haulage than Highway 128 and pro- 
vides a more convenient route to Carlsbad and 
other centers. A 24-foot (7.3-meter) wide road 
with a 4- to 6-inch ( 1 0- to 1 5-centimeter) crushed 
stone base, covered by 2 inches (5 centimeters) 
of hot-mixed asphalt without structures, is es- 
timated to cost about $40,000 per mile ($24,860 
per kilometer). 

A rail spur built into the area would probably 
be serviced by the Santa Fe Railroad as an ex- 
tension of its lines. A spur could be built as an 
extension of the Nash Draw mine or Kermac 
spurs (fig. 3). In either case, about 7 miles (11.3 
kilometers) of standard-gage rail line would be 
required. Where no major structures are needed, 
such as bridges, culverts, etc., costs would range 
from $400,000 to $500,000 per mile ($248,600 
to $310,800 per kilometer) of track. 



65 



ESTIMATES OF HYPOTHETICAL MINING UNITS IN THE WIPP SITE 



The WIPP potash deposits valuations are 
based on the economic analysis of hypothetical 
mine-mill complexes by extrapolation, where 
necessary, of the capital and operating costs of 
existing operations estimated in the previous 
section. These operations would be located 
where it is determined to be most economic and 
convenient for mining recovery and processing 
of the deposits delineated by USGS (fig. 29). 

For the purposes of this study, a deposit or 
group of deposits, suitably located for mine re- 
covery from one operation, is designated as a 
Mining Unit. Four such groups are in the site 
and were labeled A, B, C, or D. Each deposit 
size (tonnage), thickness, and lateral extent were 
estimated by USGS at three potash cutoff grades: 
high, medium, and low. The medium and low 
grade deposits include all material in the higher 
grade deposits. Each grade cutoff is considered 
a subunit, making 12 subunits — one each with 
(1) the high grade, (2) the medium grade, and 
(3) the low grade for each of the groups A— D. 
These were then designated as hypothetical op- 
erations (Mining Unit A-1, A-2, A-3, B-1, B- 
2, etc.) to evaluate the economics of potash re- 
covery in the deposit or group of deposits at 
various cutoff grades and tonnages. The bound- 
aries of the WIPP site were disregarded when 
evaluating the deposits in a Mining Unit. The 
value of only that portion of the deposit or de- 
posits within the WIPP site was then determined 
to be a loss chargeable to the WIPP facility pro- 
vided its removal did not affect the commercial 
viability of the remainder of the deposit. 

U.S. Geological Survey tonnage and grade 
data were adjusted to a minimum mining height 
of 4.5 feet (1.4 meters) and 85-percent expected 
mining extraction of the ore. Where the thick- 
ness of the mineralized zone is less than 4.5 feet 
(1.4 meters), the ore tonnage was increased by 
assuming the mineralized material would be re- 
covered in a minimum 4.5-foot ( 1 .4-meter) mine 
heading. Increasing the tonnage in this way ne- 
cessitated grade adjustments since the additional 
material is assumed to be barren of potash val- 
ues. Langbeinite grades were also adjusted due 
to an assumed dilution by 6 inches (15 centi- 
meters) of barren rock, because the undercut 



used in the mining recovery would be made be- 
low the comparatively hard langbeinite zone. 
Measured and indicated ore tonnages were 
summed and included as one value. 

Hypothetical mine-mill complexes suitable to 
the type of mineralization and desired capacity 
were developed for each Mining Unit. Results 
from the metallurgical testing were analyzed to 
determine which mineralized zones were ame- 
nable to beneficiation for each subunit. Other 
criteria, such as current and anticipated market 
demand, energy requirements and usage, and 
available water supplies, were also considered. 

The least costly method of mining and proc- 
essing was then applied to each Mining Unit. 
This required the examination of alternadve 
methods of extracting and processing each Min- 
ing Unit, with more than one system being tried. 
Consideration was given both to the economics 
of constructing new mill complexes and to uti- 
lizing portions of existing plants in the Carlsbad 
area where sulfate was one of the products. The 
capacity, using either or both of the existing sul- 
fate plants, was examined. A brief description 
of the methods, assumptions, and considera- 
tions used in evaluating each Mining Unit at the 
three grade cutoffs follows. 

MINING UNIT A 

Mining Unit A deposits consist of the western 
lobe of the 10th and 8th sylvite ore zones and 
the langbeinite of the western lobe of the 2d and 
3d ore zones. Grade and tonnage information 
for the three subunits are given in table 20. 

Mining Unit A-1 

Mining Unit A-1 consists of the western lobe 
of the 10th and 8th sylvite ore zones. No lang- 
beinite mineralization containing the high grade 
(8 percent K2O) langbeinite is in this subunit. 
Grade and impurity information is shown in ta- 
ble 21. 

As shown in table 21, the resource contains 
comparatively high grade sylvite, which could 
be beneficiated by either flotation or solution- 
crystallization methods. Kainite and carnallite 
are not present in sufficient quantity to signif- 



R30E 



R3IE 




xS-i Proposed mine location 

I I Federal surface and mineral rights 

L",^ State surface and mineral rights 

['^,1 Private surface and mineral rights 

p^ Private surface, all mineral rights owned 
"^■^ by Federal Government 

F^ Private surface and mineral rights, except 
^^^ oil and gas federally owned 

— — Proposed WIPP site outline 

— — Zone boundaries and areas provided by ERDA 



Zone 
I - 58 acres 
n -1,889 acres 
m - 6,201 acres 
nr - 10,812 acres 



FIGURE 29. — Approximate locations of Mining Units. 



67 



TABLE 20. — Average K2O grade of sylvite 

and langbeinite deposits and total tonnage, 

Mining Unit A 



Ore zone 


Weighted 
(percent) 


Total 
TMMf 


Percent 

inside 

WIPP site 


Tonnage 

inside 
WIPP site 

(MM) 


Mining Unu A- 1: 
Sylvite: 

10th, west lobe 


13.18 
13.78 


47.74 
9.86 


43 
70 


20.53 


8th _ 


6.90 






Total 


13.33 


57,60 


47.6 


27.43 






Mining Unit A-2: 
SylvTte: 


11.32 
11.24 


57.26 
41.06 


46 
62 


26.34 


8th 


25.46 






Total 


11.28 


98.32 


52.7 


51.80 






"""irJ^Uobe 


4.60 
4.76 


14.18 
26.81 


100.0 
90.0 


14.18 


3d 


24.13 








4.70 


40.99 


93.5 


38.31 






Mining Unit A-3: 
Sylvite: 

10th, west lobe 


10.65 
9.56 


69.91 
65.11 


45 
65 


31.46 


8th 


42,31 






Total 


10.03 


135.02 


54.6 


73.77 






Langbeinite: 
2d, west lobe 


4.04 
4.15 


34.40 
40.85 


50 
80 


17.20 


3d _ 


32.68 






Total 


4.12 


75.25 


66.3 


49.88 







icantly affect flotation recovery. Kieserite, while 
present, should not affect flotation. Microscopic 
tests and studies by the Salt Lake City Research 
Center indicate that liberation of the sylvite 
from the Kieserite for similar ores is possible — 
insolubles are high and would lower recovery 
and grade in a flotation circuit. To increase the 
recovery and grade, additional slimes treatment 
facilities would probably be needed to process 
this resource in the flotation circuit. In a solution 
process the insolubles would not be a major 
problem and the kieserite, at about 2.3 percent 
SO4, is within limits that can be processed com- 
mercially. 

Based on these resource and mineralization 
studies, three hypothetical mine-mill circuits 
were evaluated. 

1 . The deposit mined at a rate of 9,484 tons per 
day (8,600 metric tons per day) using a room- 
and-pillar mining method, then shipped to 
a currently operating mill for processing. 



Mining costs adjusted from the 13,000-ton- 
per-day (11,800-metric-ton-per-day) opera- 
tion to correct for the reduced mining height. 
Processing estimated by sylvite flotation re- 
covery in an existing plant. No additional 

itment circuits adc 
ery of 77 percent estimated. 

2. The same conditions as in evaluation 1 but 
with the addition of circuits for desliming and 
brine recovery; K^O recovery of 81 percent 
assumed. 

3. Analysis was made of a mine having a 7,500- 
ton-per-day (6,800-metric-ton-per-day) ca- 
pacity with the material treated for recovery 
of sylvite by solution and crystallization in an 
existing plant and with an estimated KgO re- 
covery of 81.2 percent. 

Of the three mine and mill circuits consid- 
ered, the first option, with the use of a flotation 
circuit and without additional slimes treatment, 
was found to be the most economic. 



Mining Unit A-2 

This subunit includes sylvite deposits of the 
northwest lobe of the 10th and 8th ore zones 
and langbeinite deposits from the west lobe of 
the 2d and 3d ore zones. Descriptions of re- 
source, grade, and impurities are shown in ta- 
bles 20 and 21. 

The langbeinite resource, because of its low 
grade, would be difficult to process to market 
grade, as indicated by the Salt Lake City Re- 
search Center tests. It contains more polyhalite 
and insolubles than the Salt Lake City test sam- 
ples (table 2). Calculations, assuming that the 
recovery of langbeinite and polyhalite will be 
comparable (83 percent) and with 80 percent of 
insolubles removed, show that the maximum 
concentrate will be 20.8 percent K2O (below the 
market standard of 22 percent). They further 
show that such a concentrate would contain 10. 1 
percent impurities, which is above the maximum 
allowable market amount of 4 percent. This 
langbeinite resource does not appear amenable 
to processing. For this reason, the langbeinite 
mineralized zones were not evaluated. 



TABLE 21. — Mineralogical compositions of sylvite and langbeinite deposits, Mining Unit A 

(Weight-percent) 





Mining Unit A-1 


Mining 


Unit A-2 


Mining Unit A-3 


Mineral 


Sylvite 


Sylvite 


Langbeinite 


Sylvite 


Langbeinite 


Sylvite ____ 


13.33 

3.48 
.99 
.91 

3.04 
.62 

1.19 


11.28 

5.Y9 
2.14 

.64 
2.14 

.44 
3.27 


470 
1.52 
1.46 

o' 

1.43 



10.03 

5.54 
2.29 

.57 
1.90 

.39 
3.75 




Langbeinite 


3.78 


Insolubles _ __ 


1.67 
1.80 


Kainite " 


.80 


Kieserite 

Leonite ..._ 

Carnallite 




1.43 





68 



The sylvite resource, while of lower grade 
than that in Mining Unit A-1, could be concen- 
trated by either flotation or solution-crystalli- 
zation methods. Kainite is not present in sufficient 
quantity to significantly affect flotation, and 
microscopic tests indicate the sylvite can be lib- 
erated from the kieserite. The amount of slimes 
would require modification of the previously 
discussed mill operations for extra slimes treat- 
ment. The amount of carnallite is only slightly 
greater than 2 percent and could be tolerated 
in a solution-crystallization circuit with an esti- 
mated KjO recovery slightly lower than that ex- 
pected for Unit A-1; carnallite may also affect 
the recovery in a flotation circuit. 

Based on these assumptions, sylvite resources 
in Mining Unit A-2 were evaluated for two min- 
ing and milling circuits. 

1 . The deposit would be mined at a rate of 9,484 
tons per day (8,600 metric tons per day) using 
a room-ancl-pillar mining method. The ma- 
terial is then shipped to an existing mill based 
on flotation recovery. The existing plant 
would be modified for extra desliming and 
brine recovery circuits. A KoO recovery of 80 
percent is estimated. 

2. The deposit would be mined at a rate of 7,500 
tons per day (6,800 metric tons per day). The 
material would be recovered by solution-crys- 
tallization methods from an existing plant. 
The plant would be modified to include a 
carnallite circuit. A recovery of 80.0 percent 
is assumed. 

The mine and flotation complex of evaluation 
1, as modified with additional desliming and 
brine recovery, was the more economic. 



value of 4 percent. Thus, the langbeinite re- 
source does not appear amenable to processing. 
The sylvite resource, while of lower grade than 
that in Mining Unit A-1, could be concentrated 
by either flotation or solution-crystallization 
methods. Kainite is not present in sufficient 
quantity to significantly affect flotation, and 
kieserite should not affect flotation appreciably 
since microscopic tests indicate the sylvite can 
be liberated. The amount of slimes is high 
enough to necessitate modification of mill models 
for extra slimes treatment. The carnallite may 
affect recovery in a flotation or solution-crys- 
tallization plant. 

Two mine and mill operations were proposed. 
The proposed mine for the extraction of the 
sylvite mineralized zones would be identical to 
that in the previous section on Mining Unit A- 
1 but with extended project life because of the 
larger resource amount. 

The related mill circuits would be: 

1. Sylvite would be recovered by flotation 
through the flotation circuit of an existing 
plant, modified with additional desliming 
and brine recovery circuits, and would pro- 
vide an estimated K2O recovery of 80 per- 
cent. 

2. Sylvite would be recovered by a solution 
and crystallization method in an existing 
operation and would provide an estimatecl 
K2O recovery of 80 percent. The plant 
would be modified to include a carnallite 
circuit. 

A flotation circuit, modified with additional 
desliming and brine recovery, was determined 
to be the more economic milling circuit. 



Mining Unit A-3 

Mining Unit A— 3 is a combination of sylvite 
deposits in the 8th ore zone and the western 
lobe of the 10th ore zone and langbeinite de- 
posits occurring in the western lobe of the 2d 
and 3d ore zones. These mineralized zones in- 
clude the zones mentioned previously in Mining 
Units A-1 and A-2, plus additional lower grade 
material. Resource and impurities within these 
mineraUzed zones are shown in tables 20 and 
21. 

The langbeinite resource is low grade and 
contains significant amounts of polyhalite and 
insolubles. Calculations, assuming a recovery of 
langbeinite and polyhalite of 83 percent and 90 
percent removal of insolubles, respectively, in- 
dicate that the concentrate would have a grade 
of 19.9 percent K2O equivalent as langbeinite 
and 12 percent impurities. This amount of im- 
purities is above the maximum allowable market 



MINING UNIT B 

The hypothetical Mining Unit B was designed 
to extract and process material from the 4th and 
5th langbeinite-mineralized ore zones. Follow- 
ing is a brief description of the methods and 
assumptions used in the evaluation of hypo- 
thetical Mining Unit B, 

Mining Unit B-1 

Mining Unit B-1 (highest grade) includes 
langbeinite deposits of the 4th and 5th ore 
zones. Average mining grade of langbeinite 
(KgO equivalent) and reserve within the WIPP 
site are given in table 22. 

A hypothetical 6,000-ton-per-day (5,400-met- 
ric-ton-per-day) capacity langbeinite leach plant, 
constructed at the site, was assumed. In addition 
to a langbeinite product, this plant would pro- 
duce enough langbeinite fines to supply sulfate 
sections at two existing plants. A conventional 



69 



TABLE 22.— Average KgO ^ade of 
langbeinite and total tonnage, Mining Unit B 



Ore zone 


Weighted 

grade 
(percent) 


Total 
"(MMf 


Percent 

inside 

WIPP site 


Tonnage 

inside 
WIPP site 

(MM) 


Mining Unit B-1: 

4th*. 

5th 


9.18 
7.20 


72.23 
7.55 


65 
20 


46.95 
1.51 


Total 


9.11 


79.78 


60.8 


48.46 


Mining Unit B-2: 


7.69 
4.59 


126.33 
27.85 


57 
85 


72.01 


5th 


23.67 


Total 


6.92 


154.18 


62.1 


95.68 


Mining Unit B-3: 
4th north 


6.87 
4.02 


164.83 
30.23 


60 
82 


98.9U 


5th 


24.89 


Total ___. 


6.30 


196.07 


63.1 


123.79 



mining system with adjusted costs to mine a 4.5- 
foot (1.4-meter) bed of langbeinite was used (see 
appendix for further detail). All costs of mining 
sylvite needed for sulfate production were added 
to the costs of sulfate production. 

The amount of polyhalite, insolubles, and 
kieserite present in the resource (table 23) are 
within current processing tolerances. Calcula- 
tions indicate that a concentrate grade of 21.6 
percent K.^O equivalent or better could be made 
including 4 to 4.5 percent impurities. These cal- 
culations assume a comparable recovery of po- 
lyhalite and langbeinite and 80 percent removal 
of insolubles. These grade and impurity content 
calculations indicate a concentrate meeting mar- 
ket specifications could be recovered. This re- 
source is leachable and the concentrate fines 
could then be combined with purchased sylvite 
fines to make a sulfate product. 

Mining Units B-2 and B-3 

Mining Units B-2 (medium grade) and B-3 
(lowest grade) include langbeinite deposits from 
the 5th ore zone and the north portion of the 
4th ore zone. Average mining grades of lang- 
beinite (KjO equivalent) and total resources in 
the WIPP site are given in table 22. 

Langbeinite resources in both Units B-2 and 
B-3 (table 22) are low grade and contain amounts 
of high impurities so that, by calculation, neither 



TABLE 23. — Mineralogical compositions of 
langbeinite deposits, Mining Unit B 



(Weight-percent) 






Mineral, average comp. 


Mining Unit 


Mini„g^Un.t 


M.„.ng^Unit 


Langbeinite 


9.11 
2.07 
1.38 
l.IO 
.47 
.93 
1.17 


6.92 
3.15 
1.41 
.98 
.66 
1.05 
2.23 


6 30 


Insolubles 

Polyhalite _ 

Kainite 

Kieserite 

Leonite 

Carnallite 


3.21 
1.42 
1.01 
.68 
1.00 
191 







would meet market grades through current 
commercial processing. The calculation assumes 
comparable recoveries of langbeinite and poly- 
halite in the concentrate and 80 percent of the 
insolubles removed. For Unit B-2, calculations 
indicate the best product grade that could be 
expected would be about 21 percent K^O with 
7 percent impurities, and for Unit B-3 about 
20.9 percent K^O with 7.8 percent impurities. 

These calculations indicate the langbeinite re- 
sources in both Units B-2 and B-3 are not ame- 
nable to commercial processing, and no economic 
evaluations were performed in either case. 

MINING UNIT C 

Mining Unit C deposits consist of the eastern 
lobe of the 10th ore zone, the 9th sylvite-mi- 
neralized ore zone, and the east lobe of the 2d 
and 10th ore zones. Grade and tonnage infor- 
mation on resources for the three subunits are 
given in table 24. 

Mining Unit C-1 

Mining Unit C-1 includes the 10th and 9th 
mineralized zones, which contain sylvite miner- 



TABLE 24.— Average KjO grade of sylvite 

and langbeinite deposits and total tonnage. 

Mining Unit C 



Ore zone 


Weighted 

^graT 
(percent) 


Total 
"MMf 


Percent 

inside 

WIPP site 


Tonnage 

inside 
WIPP site 

(MM) 


Mining Unit C-I: 
Sylvite: 
10th 


14.51 
12.65 


31.20 
.64 


50 
100 


15 60 


9th 








Total 


14.44 


31.84 


51 


16 24 






Langbeinite: 
lOth 


8.12 
8.06 


6.90 
11.60 


100 

78 


6 90 


2d 








Total 


8.08 


18.50 


86.2 


15 95 






Mining Unit C-2: 
Sylvite: 

lOth east lobe 

9th.. 


11.57 
10.16 


51.75 
5.44 


60 
100 


31.05 






Total 


11.36 


57.19 


63.8 








Ungbeinite: 

icfth east lobe 

2d east lobe 


4.91 
5.49 


50.74 
41.81 


86 
90 


43.64 
37.63 


Toul 


5.18 


92.55 


87.8 


81.27 


Mining Unit C-3: 
Sylvite: 

10th east lobe 

9th 


10.59 
9.56 


61.28 
9.36 


71 
100 


43.51 
9.36 


Total _. 


10.41 


70.64 


74.8 


52.87 


Langbeinite: 

10th east lobe 

2d east lobe 


4.38 
5.13 


52.03 
52.53 


95 
95 


49.43 
49.90 


Total 


4.76 


104.56 


95 


99.34 



70 



TABLE 25. — Mineralogical compositions of sylvite and langbeinite deposits, Mining Unit C 

(Weight-percent) 





Mining Unit C-l 


Mining Unit C-2 


Mining Unit C-3 


Mineral, average composition 


Sylvite 


Langbeinite 


Sylvite 


Langbeinite 


Sylvite 


Ungbeinite 


SyUite 

li^rotbier 

Polyhalite 

Kain.te 

Kieserite 

Leonite 

Carnall.te 


14.44 

4.7 
2.06 

9.02 

".39 


2.59 
1.64 
1.51 


11.36 

3.98 
1.9 

6.52 


2.40 
1.55 
1.71 
.55 
1.45 


10.41 

3,82 
1.81 

6.25 


476 
2.40 
1.52 
1.61 
.49 
1.45 



alization. In addition, tliis subunit contains lang- 
beinite mineralization in the 10th and 2d ore 
zones. 

Analysis of the langbeinite resource in Unit 
C— 1 indicates that the deposits are not amenable 
to current commercial processing to market 
standards. The optimum product would contain 
20.5 percent K„0 and 9.6 percent impurities. 
For these reasons, the resource is not amenable 
to commercial processing. 

The sylvite from Unit C-l would be recover- 
able by flotation. The amount of kieserite in the 
resource (table 25) contains enough SO^ (6 per- 
cent) to cause difficulties in a crystallization cir- 
cuit but would not affect a flotation circuit. 
Bureau of Mines tests indicate the sylvite would 
be liberated. The insolubles content of the re- 
source is sufficiently high to require extra de- 
sliming and brine recovery circuiting. 

The mining system used for this hypothetical 
subunit has a design capacity of 3,720 tons per 
day (3,400 metric tons per day) resulting in a 
project life of 15 years. The costs estimated for 
this mining system were adjusted from the syl- 
vite-mine portion of the 7,000-ton-per-day 
(6,398-metric-ton-per-day) operation discussed 
in a previous section. The mined material would 
be shipped to an existing mill for processing to 
a muriate product. 

Mining Unit C-2 

This subunit includes sylvite products in the 
9th ore zone and the eastern portion of the 10th 
ore zone, and langbeinite deposits in the eastern 
lobe of the 2d and 9th ore zones. Average min- 
ing grade (KgO equivalent) is 11.36 percent of 
sylvite and 5.18 percent for langbeinite (table 
24). Analysis of the impurities in the langbeinite 
(table 25) indicate the potash product would not 
meet market standards. A 20.6 percent K.^O 
grade product containing 8.8 percent impurities 
is estimated. For these reasons, the langbeinite 
mineralized zones were not evaluated. 

The sylvite deposit in Unit C-2 is estimated 
to be processable by flotation. The amount of 
kieserite is at the high (2.99 percent SO4) limits 
for a standard solution circuit, but microscopic 



tests indicate that the sylvite can be liberated 
from the kieserite, which is necessary for flo- 
tation separation. The insolubles content is high 
enough to require extra desliming and brine 
recovery circuiting over that designed in the 
operations described previously. 

In evaluating Mining Units C-2 and C-3, a 
larger mining capacity [9,486-ton-per-day (8,- 
604-metric-ton-per-day) operation] was esti- 
mated for the sylvite ore zones because of the 
larger resource tonnage at the medium and low 
grade levels. Evaluation at this increased capac- 
ity resulted in lower cost per unit. A sylvite flo- 
tation mill was modified to include brine recovery 
and extra desliming, resulting in a K.,0 recovery 
of 80 percent. 

Mining Unit C-2 was found to be slightly 
more profitable than Mining Unit C-I (the 
higher grade) because economy of scale in the 
mine and the mill in this case results from the 
capacity increase. For this reason, total resource 
revenue on subunits C-l and C-2 is equal to the 
revenue from the medium grade deposit unit 
(Mining Unit C-2). 

Mining Unit C-3 

Mining Unit C-3 (lowest grade) includes syl- 
vite deposits in the 9th ore zone and the eastern 
lobe of the 10th ore zone and langbeinite de- 
posits in the eastern lobes of the 2d and 3d ore 
zones (table 24). Average mining grade (K^O 
equivalent) was 10.41 percent for sylvite and 
4.76 percent for langbeinite. Analysis of the im- 
purities in the langbeinite deposits indicates that 
it would yield a product containing 20.5 percent 
KjO grade with 9.6 percent impurities that does 
not meet market standards (table 25). For this 
reason, the langbeinite mineralization was not 
evaluated. 

The sylvite resource from this unit could be 
beneficiated by flotation. Kieserite should not 
preclude flotation recovery, since the sylvite can 
be liberated, but the SO4 content is too high 
(2.99 percent) for recovery by solution-crystal- 
lization. The insolubles are in sufficiently high 
concentrations to require extra desliming and 
brine recovery circuiting in the flotation plant. 



71 



Based on these assumptions, the resource was 
evaluated for sylvite flotation, modified to in- 
clude extra desliming and brine recovery cir- 
cuits, with estimated K^O recovery of 78 percent. 
A conventional room-and-pillar mining sys- 
tem, using the same mining capacity as Mining 
Unit C-2, was used to extract this low-grade syl- 
vite mineralization. 

MINING UNIT D 

This Mining Unit was used to evaluate the 
southern portion of the 4th langbeinite ore 
zone. No evaluation was performed on the high- 
est grade resource (Mining Unit D-1) because 
sufficient tonnage to be commercial is not pre- 
sent. Evaluations were made on the medium and 
low grade resources. Resource data are shown 
in table 26. 

Mining Unit D-2 

Mining Unit D-2 contains langbeinite miner- 
alization from the southern portion of the 4th 
ore zone. As shown in table 26, the medium 
grade resource has an average mining grade of 
5.00 percent (K.^O equivalent). 

The resource of Mining Unit D-2 contains 
only a small amount of impurities, as illustrated 
in table 27. Because the percentage of impurities 
and insolubles is below that of synthetic low- 
grade samples analyzed by the Salt Lake City 
Research Center, the deposits could be proc- 
essed commercially to yield a marketable prod- 
uct. 

TABLE 26.— Average KgO ^rade of 
langbeinite and total tonnage, Mining Unit D 

(Weight-percent) 



Ore zone 


Weighted 

grade 
(pecent) 


Total 
tonnage 


Percent 

inside 

WIPP site 


Tonnage 
inside 

WIPP site 
(MM) 


Mining Unit D-2: 

4th 

Mining Unit D-3: 

4th 


5.00 
4.30 


87.93 
140.27 


26.8 
30.3 


23.57 
42.45 



TABLE 27. — Mineralogical compositions of 
langbeinite deposits, Mining Unit D 



(Weight-percent) 




Mineral, 
average composition 


Mining Unit D-2 


Mining Unit D-3 


Langbeinite 


5.00 
1.57 
.04 
.09 
.004 


4.30 


Insolubles 


1.57 


Polyhalite 


.03 


KaiW 

Leonite 


.08 
.03 



A conventional room-and-pillar mining sys- 
tem with a capacity of 10,000 tons per day (9,070 
metric tons per day) was used in this evaluation. 
The mined material would be sent to a lang- 
beinite wash plant built at the site. A langbeinite 
product would be produced, and the resulting 
fines would be shipped and processed at two 
existing sulfate plants. 

All costs of mining a sufficient amount of syl- 
vite to produce the necessary amount of fines 
used with the langbeinite fines in the sulfate 
process were charged to the mining cost of this 
subunit. 

Mining Unit D-3 

This subunit consists of the lowest grade re- 
source found in the 4th ore zone. The average 
mining grade and resource data are shown in 
table 26. The impurities are low, indicating that 
the langbeinite could be processed to a market- 
able product. 

The same type of mine-mill complex de- 
scribed in D-2 was used in the evaluation of 
Mining Unit D— 3 but at a reduced capacity of 
7,000 tons per day (6,350 metric tons per day) 
in order to provide sufficient langbeinite fines 
for only one sulfate plant. A langbeinite wash 
plant would be built at the site. The langbeinite 
fines from this wash plant would be shipped to 
an existing sulfate plant for further processing. 
All costs associated with the mining and milling 
of sylvite to produce the necessary fines re- 
quired in the production of a sulfate product 
were charged to the D-3 mine and mill complex. 

ESTIMATED MINING UNITS CAPITAL 
INVESTMENTS AND OPERATING COSTS 

Table 28 shows operating costs and capital 
investments for 8 of the 12 hypothetical Mining 
Units. These Units were selected for analysis to 
determine if they were commercial at 1977 mar- 
ket product prices and estimated costs. No de- 
posits corresponding to Mining Unit D-1 (highest 
grade) are in the site; Mining Units B-2 and B- 
3 were eliminated because they contain material 
that could not be treated by current technology 
to recover marketable products; Mining Unit C- 
1 mineralization was included with that for Unit 
C-2 (because of better economics, it would be 
mined and processed as that unit). Table 29 
shows estimates of product recovery as sylvite 
(or muriate), langbeinite, and sulfate. 



72 



TABLE 28. — Summary of operating costs and capital investments for hypothetical Mining 

Units 





Mining 
Unit"* 


Operating costs', $/st 


Capiul investment, $1,000 




Mine 


Mill 


Freight 


Mine 


Mill 


Toul 




3.55 
3.55 
3.55 
5.64 
3.55 
3.55 
4.58 
4.89 


2.67 
2.67 
2.67 
7.24 
2.67 
2.67 
4.34 
4.56 


2 54 


$35,652 
35,652 
35,652 
56,042 
35,652 
35,652 
37,439 
31,256 


$ 3,530 
3,530 
3,530 
8,447 
3,530 
3,530 
11,826 
10,213 


$39,182 


A-2 


2 
2 


54 
54 
04 
54 
54 
02 
02 


39,182 
39,182 


B-1 


64,489 


C-2 . 
C-3 . 
I>-2 . 
D-3 _ 




39,18$ 
39.182 
49.265 
41.469 



' Costs of obtaining sufficient sylvite (where sylv 
langbeinite mining and milling. 
J Added to mine operating costs. 
Added to mill operating costs. 



not mined at the operation) to process langbeinite fines for sulfate produ 



73 



m 



£2l5' 



.SI 



mu 



i i 



s. g = " s. s 



lis 



iii I til ^. til til til ^- til i til 



! Ij 1 ! 
1 li I • 

i I III " H 






i« S«5 Sf Stl 



i i i 



74 



SUMMARY AND CONCLUSIONS 



The Bureau of Mines evaluation of potash 
deposits in the WIPP site and subsequent eco- 
nomic and financial analyses of hypothetical 
mining and processing operations show that one 
of the 12 Mining Units — Unit B-1, parts of 
which are in the site — is commercial at current 
mining and processing costs and at present mar- 
ket prices. These deposits are thus ore and are 
classified as reserves. 

Potash deposits included in these Mining 
Units occur in various ore zones accessible from 
proposed shaft locations. The mining and be- 
neficiation systems evaluated are based on cur- 
rent extraction and processing technology in the 
district and are the least costly systems amenable 
to the ore mined. An economic evaluation of the 
mineralization determined to be in the site near 
its western boundary was made considering 
mining recovery through an existing shaft in 
addition to recovery through a new shaft. The 
grade of the mineralization and deleterious min- 
eral content determined from drill core samples 
indicate that this mineralization is not commer- 
cial using current Carlsbad technology. 

This analysis isolates and estimates the values 
that exist in the unmined potash mineralization 
in the WIPP site and, therefore, determines a 
cost chargeable to the WIPP facility. These val- 
ues are items of costs and capital investments 
(taxes, royalties, and bonus bid amounts) that 
would be generated from potash product sales 
and would be paid to the several levels of gov- 
ernment. They are losses because the tax and 
royalty revenues and bonus bid amounts, or 
their equivalent value, will not accrue to the gov- 
ernments if the potash deposits are not mined 
and produced. Unlike taxes, royalties, and bo- 
nus bid amounts, all other items of production 
costs, investments, and values are created when 
the projects are initiated (when capital is in- 
vested) and, therefore, are not a loss if the un- 
mined, commercial potash is not developed. 

Loss to the owners (other than government) 
would be royalty payments and a value deter- 
mined for the unmined potash (equivalent to 
the bonus bid amount due the governments). 
The taxes that would have been generated if the 
potash were produced from privately owned 



potash would be a loss to the governments. 
These values (taxes and royalties — because they 
are estimated to occur in a series of lump sums 
over the life of the project — and the bonus bid 
amounts) are converted (discounted at an ap- 
propriate discount rate) to equivalent values at 
a single point in time (usually project initiation). 
In this analysis, it results in an amount in current 
dollars or the present value. The present or cur- 
rent value (as commonly used in economic anal- 
yses) of the unmined potash (bonus bid amount) 
is the dollar amount that the gross revenues ex- 
ceed the sum of the (1) costs of production, (2) 
return of the investment, and (3) interest on that 
investment; all items discounted at a future 
worth factor of money. The present value is the 
amount that an investor would be able to pay 
for the unmined potash and still receive (1) the 
return of his total investment and (2) an ac- 
ceptable rate of interest on his total investment 
including the amount he pays for the unmined 
potash. In this analysis, for those government- 
owned deposits that are commercial, the present 
value amount is added to the investments as a 
lease bonus bid (a dollar amount paid as a bonus 
to obtain a lease). 

Items such as the return of the original in- 
vestment and interest on that investment as well 
as all direct and indirect costs of production 
must be provided from gross revenue in a prof- 
itable operation. This can be expressed as fol- 
lows: The gross revenue from the sale of all 
potash products equals all direct and indirect 
costs including (1) materials, (2) labor, (3) sup- 
plies, (4) utilities, (5) taxes, (6) royalties, and (7) 
other costs plus the interest on the investment 
plus return of the investment for (1) explora- 
tion, (2) development, (3) mine plant, (4) mill 
plant, (5) bonus bid amount, and (6) all other 
capital required. 

The major portion of the dollar amounts of 
gross revenues is used to purchase goods and 
services required for potash production. 

One group of the highest grade deposits, des- 
ignated as Mining Unit B-1 and occurring par- 
tially in the site, was determined to be commercial. 
About 20.2 million tons (about 18.3 million met- 
ric tons) of potash products could be produced 



75 



140 



126 



112 



:^ 98 



CTi 

- 84 



o 

=3 
O 

^ 70 



56 



y 42 



28 



14 



Prices include 15-percent return 
on invested capital 



Sulfate 



Sylvite 



r" 



Langbeinite 



6 9 12 15 

PRODUCT, million short tons 



18 



21 



FIGURE 30. — Product availability and market prices at which potash tonnages become commercial 
assuming fixed production costs. 



76 



90 



81 



72 



63 



54 



45 



36 



27 



18 



\ / y/\ Mine operating cost 

^^^^ Mill operating cost 

|^\\^^ Tax and royalty costs 

I I R^ 

I 1 in 



Return of investment and 
terest costs 




7 14 21 28 35 42 

ADDITIONAL POTASH PRODUCTS, million short tons 

FIGURE 31. — Approximate cost analysis of increased market prices at which potash tonnages 

become commercial. 



77 



from the 48.46 million tons (43.95 million metric 
tons) of ore in these reserves. The value lost 
(dollars that would not be generated from po- 
tash production) in terms of current dollars for 
an acquisition cost, for taxes payable on esti- 
mated revenues, and for royalties payable to the 
owners of the potash, is about $51.8 million. The 
gross market value of the products is about 
$1.03 billion which, less operating costs other 
than taxes and royalties, less the return of the 
investment other than the acquisition cost, and 
less interest on the investment limited to 15 per- 
cent, is the $5 1 .8 million value chargeable to the 
WIPP project. The procedure used to calculate 
an estimate of the values that will be foregone 
if the WIPP installation goes forward is shown 
in table 30. 

The amounts of potash products that could 
be recovered from potash reserves and re- 
sources in the WIPP site, at specific market 
prices and associated costs, are illustrated in fig- 
ures 30 and 31 and in tables 31, 32, and 33. 
Evaluation data are summarized in table 29. 

An additional 169.1 million tons (153.4 mil- 
lion metric tons) of lower grade, paramarginal 
potash resources in the site contain 25.5 million 
tons (23.1 million metric tons) of products if 
today's operating costs did not increase and 
market prices were increased until the deposits 
become commercial. The dollar values of these 
subeconomic resources are speculative; how- 
ever, if the relationship of actual losses to gross 
market values exists as in the commercial de- 
posits, the value lost at some future time when 



they become commercial is about $75.4 million 
from a gross market value of products of about 
$1.5 billion. Recent market prices are shown in 
table 34. 

Products that meet market standards cannot 
be produced, using current Carlsbad processing 
techniques, from the remainder of the miner- 
alization determined to be in the site. This mi- 
neralization is considered to be a submarginal 
resource. 

An option is being considered that would al- 
low commercial recovery of potash products in 
WIPP Zone IV. Zone IV is a comparatively large 
area — 10,812 acres (4,376 hectares) that con- 
tains much of the mineralization in the WIPP 
site. For this reason, the estimated amounts and 
values of potash product that would be foregone 
if mining operations were prohibited only in 
Zones I, II, and III are 13.3 million tons (12.1 
million metric tons) of ore (commercial miner- 
alization) containing 5.5 million tons (5.0 million 
metric tons) of potash products (table 31). The 
value lost in terms of current dollars is about 
$14.3 million. The gross market value of the 
products is about $282.4 million. The gross mar- 
ket value of the products is about $282.4 million. 
Paramarginal resources in 51.37 million tons 
(46.59 million metric tons) of potash minerali- 
zation contain 7,640,500 tons (6,931,400 metric 
tons) of potash products. The ratio (the same as 
for the commercial resources) of value lost to 
gross product market value is about $17.5 mil- 
lion from a gross market value of $346.3 million. 



TABLE 30. — Mining Unit B-1 estimated potash values within the WIPP site that would be 

foregone 



Tonnage: 

Total in deposit 

Total in Mining Unit B-1 operation' 

Part of Mining Unit B-1 in WIPP site 

Total of B-1 in WIPP site on State lands (all in Sec 16, 
Total of B-1 in WIPP site on Federal lands 



1,000 short Im 
of ore 
79.780 
65,250 
48,460 

7,753.6 
40,706.4 





Mining Unit 

at 8 percent 

over 50 years Va 
($1,000) J 


lue, 

/st 


WIPP 
Taof 


Revenues lost ($1,000) 


Item 


State 
Govt. 


Federal 
Govt. 


"^'t^;^ --:-::-:::: 


19,021 

2;383 

5,060 
19,632 
23,325 


29 
29 
29 
04 
08 
30 
36 
36 
36 


40'706.4 
7,753.6 
48,460 
48,600 
48,460 

40706.4 
7,753.6 


5^962 
2,249 
1,938 
3,877 

2791 


5"962 


State 

State income taxes 

Severance taxes' 

Federal income taxes 


14"5"38 






federal::::::::::::::::::::::::::::::: 

State 


14,654 


Total 


69,421 




__._ 


16,757 


35,094 




$51 


851 







the 



s was calculated l)v iiuiltiplviiiL' anruial piodu 
assigned to total tonnage in llic deposit, 
ederal royalties are divided 50-50 Iwtween the l-ederal and Stale governments, 
axes include severance, property, and priKessors ia.\es. 



1 used to derive 



78 



TABLE 31. — Summary of amounts and values of potash mineralization in WIPP site 





Products 


1.000 short tons 




Value, 


$1,000 




Mining 
Unil 


Recoverable 


Recoverable 

ore in 

WIPP site 


Product 
inW^Pp 


Price" 

that 

provtdes 


Gross 
revenue of 
products 
inside the 
WIPP site 


Gross' 
revePieof 
addiuonal 
products 


'v^l^e"' 

foregone 

in WIPP 

site 


B-1 


Langbeinite 

Sullate 


79,780 

57',600 
87,930 

98,320 
57.190 
140,270 

70,640 
135,020 


48,460 

27;4i0 
23,570 

5r86o 
36.490 
42.450 

52370 
73,770 


14.175 
5,998 
4.936 
3,769 
1.598 
7.791 
5,527 
5,724 
2,767 
7,293 
9,743 


35.00 
88.50 
52.04 
36.51 
92.31 
61.73 
61.74 
42.26 
106.86 
67.52 
70.28 


496,125 
530,823 
256,869 


1,026,948 

256,869 
137,606 
147,511 
176.239 
341.237 
82.618 
124,919 
119,240 
137.165 


51.851 


D-2 

.0 


Langbeinite .._. 

Sulfate 

Muriate 


:::: 








D-3 
C 3 


Langbeinite 

Sulfate 

Muriate 




A-3 


Muriate 





Market price 



produ 



Estimated weighted 



f.o.b.. Garlsbad, N. Mex.. used for 



' Product amounts that would be recovered 
lower grade Mining Units (A-2. etc.). 



ligher grade Mining Units (A-l. B-1. C-1) are subtracted from product : 



uld be recovered i 



Present values at Increased potash products market prices of resources foregone in WIPP site were not estimated because the time and conditions at which they 
Lild become commercial are too speculative. For this study, the 1975-1976 weighted average prices received for potash products in the Carlsbad area were used 
estimate the values of commercial potash deposits (ore). 



TABLE 32. — Mining Unit product data and required market prices at which potash tonnages 
in the WIPP site become commercial at fixed production costs 



Market price a 







(1.000 short 


ons) 












Price' that 


Total 


Total 


Net 








provides a 

15 percent 

ifoR^ 


Mining Unil 
product 
tonnage 


product 


additional 


Cumulative 


Mining 
Unit 


Products 


tonnage 
in WIPP 


product 
tonnage 


product 
tonnage 


B-1' 


L.ng^..te .................... ...... 


$35.00 


19,086 


14,175 


14,175 


14,175 




88.50 


8,077 


5.998 


5,998 


20,173 


A-l 


Muriate 


52.04 


9,967 


4.936 


4,936 


25,109 


D-2 


Langbemitc 
Sulfate 


36.51 


12,840 


3,769 


3,769 


28,878 




92.31 


5,446 


1,598 


1,598 


30,476 


A-2 


Muriate . 


61,73 


14,578 


7,791 


2,855 


33,331 


C-2 


Muriate 


61.74 


8,914 


5,527 


5,527 


38,858 


D-3 


Langbeinite 

Sulfite 


42.26 


11,783 


5,724 


1,955 


40,813 




106.86 


5,697 


2,767 


1,169 


41,982 


C-3 


Muriate 


67.52 


9,550 


7,293 


1,766 




A-3 




70.28 


13,715 


9,742 


1,952 














For this study the 1975-1976 weightecl-average prices received for potash prodi 

deposits. 

^ Rate of return. 

' Commercial at recent weighted-average annual prices. 



costs. Estimated weighted average annual price per ton of product, f.o.b., Carlsbad, New Mex., used forevaluation. 
'the Carlsbad area were used to estimate the values of commercial potash 



TABLE 33. — Mining Unit price components per ton of weighted-average product price 
assuming fixed production costs and a 15-percent discounted cash flow rate of return 



Mining Unit 


B-1 


A-l 


D-2 


A-2 


C-2 


D-3 


C-3 


A-3 


Product, short tons: 

Annual _ 

Additional within WIPP _ 


905,400 
20,173,100 


622,900 
4,935,900 


831,100 
5,367,300 


520,600 
2,854.600 


524,300 
5,526,900 


582,700 
3,124,300 


477,500 
1.765,900 


457.200 
1,951,700 


Cost per ton product: 


$13.55 
17.48 
6.80 

13.08 


$22.71 
14.84 
3.60 

10.89 


$20.11 
19.15 
4.65 

9.22 


$27.18 
17.75 
16,56 

10.24 


$26.98 
17.63 
5.15 

11.98 


$24.44 
22.89 
5.65 

10.33 


$29.63 
19.36 

5.85 

12.68 


$30.94 


Mill operating cost 


20.22 


Taxes and royalties 

Interest and amortization of 


6.46 
12.66 






Total _-_ 


52.04 


52.04 


53.13 


61.73 


61.74 


63.31 


67.52 


70.28 



TABLE 34. — U.S. average market prices for potash products — f.o.b. plant' 

(Dollars per short ton) 





1973 


1974 


1975 


1976 


Product 


1st 6 mo. 


2d 6 mo. 


1st 6 mo. 


2d 6 mo. 


Muriate, all grades, at 61 percent KjO: 

Per tonko ...._ .___ 


36.39 
22.20 

82.68 
42.17 

88.41 
19.45 


52.50 
32.03 

110.79 
56.50 

119.41 
26.27 


75.19 
45.87 

163.39 
83.33 

149.73 
32.94 


76.56 
46.70 

181.63 
92.63 

155.73 
34.26 


70.91 
43.26 

177.00 
90.27 

149.27 
32.84 


59.81 
36.48 


Sulfate, at 51 percent K2O: 


224.02 




114.25 


Langbeinite, at 22 percent KjO: 


154.77 


Per ton of product _ — 


34.05 



' Product market prices vary considerably from month to month and from company to company. The potash market and product prices have been seasonal and, 
in addition to the usual market supply-demand fluctuations, have been subject to discounting, inventories, buildup pressures, and product grade variations. 



79 



REFERENCES 



1. Bachman, G. D. Surfical Features and Late Cenozoic 

History in Southeastern New Mexico. U. S. (ieol. Sur- 
vey Open-File Rept. 4339-8, 1973, 32 pp., illus., maps. 

2. Brokaw, A. L.. C. L. Jones, M. E. Clooley, and W. H. 

Hayes. Geology and Hydrology of the Carlsbad Potash 
Area, Eddv and Lea Counties, N. Mex. U. S. Geol. 
Survey Open-File Rept. 4339-1, 1972, 86 pp. 

3. Coons, A, T. Potash: Mineral Resources of the United 

States, 1930, BuMines, pt. 2, 1932, pp. 59-67. 

4. Cooper, J. B., and V. M. Glanzman. Geohydrology of 

Project Gnome Site, Eddy County, N. Mex. U. S. Geol. 
Survey Prof. Paper 712-^A, 1971, 24 pp. 

5. Eilertsen, D. E. Potash. BuMines Minerals Yearbook 

1968. V, 1-2, 1969, pp. 943-944. 

6. Gale, H. S. Potash. Mineral Resources of the United 

States, 1916, U. S. Geol. Survey, pt. 2, 1919, pp.74, 
104. 

7. Gale, H. S., and W. B. Hicks. Potash: Mineral Resources 

of the United States, 1917, U. S. Geol. Survey, pt. 2, 
1920, pp. 425-426. 

8. Gard, L. M., Jr. Geologic Studies, Project Gnome, Eddy 

County, N. Mex. U. S. Geol. Survey Prof. Paper 589, 
1968, 33 pp. 

9. International Minerals and Chemical Corp. Potash, the 

Mineral With a Future. Northbrook, 111., 1968, 28 pp. 

10. Jones. C. L. Potash Resources in Part of Los Medanos 

Area of Eddy and Lea Counties, N. Mex. U. S. Geol. 
Survey Open-File Rept. 75-407, 175, 37 pp. 

11. Jones. C. L., M. E. Cooley, and G. D. Bachman. Salt 

Deposits of the Los Medanos Area, Eddy and Lea 
Counties, N. Mex. U. S. Geol. Survey Open-File Rept. 
4339-7, 1973, 67 pp. 

12. Jones, C. L,, and B. M. Madsen. Evaporite Geology of 

the Fifth Ore Zone, Carlsbad District, Southeastern 
New Mexico. U. S. Geol. Survey Bull. 1252-B, 1968, 
21 pp. 



13. Jones, C. L., C. G. Bowles, and K. G. Bell. Experimental 

Drill Hole Lagging in Potash Deposits of the Carlsbad 
District. N. Mex. U. S. Geol. Survey Open-File Rept., 
I960, 25 pp. 

14. Keyes, W. F. Potash. Ch. in Mineral Facts and Problems. 

BuMines Bull. 667, 1975, pp. 855-869. 

15. . Potash. BuMines Minerals Yearbook 1974, v. I, 

1975, pp. 1097-1109. 

16. . Potash. BuMines Commodity Data Summaries, 

Washington, D. C, 1976, pp. 128-129. 

17. Koepke, W. E. Structure, Behavior, and Performance 

of the World Potash Industry. Mineral Bull. MR 139, 
Mineral Development Sector, Dept. of Energy, Mines 
and Resources, Ottawa, Canada, 1973, 98 pp. 

18. Lewis, R. W. Potassium. Ch. in Mineral Facts and Prob- 

lems. BuMines Bull. 650, 1970, p. 1158. 

19. Nourse, M. R. Potash. Mineral Resources of the United 

States, 1920. U.S. Geol. Survey, pt. 2, 1923, p. 109. 

20. . Potash. Mineral Resources of the United States, 

1921, U.S. Geol. Survey, pt. 2, p. 54. 

21. Silver, B. A., and R. G. Todd. Permian Cycle Strata, 

Northern Midland and Delaware Basin, West Texas 
and Southeastern New Mexico, Bull. AAP(i, v. 53, 
No. 11, 1969, pp. 2223-2251. 

22. Sondermeyer, R. V. The Mineral Industry of New Mex- 

ico. BuMines Minerals Yearbook 1972, v. 2, 1974, p. 
488. 

23. Stermole, F. |. Economic Evaluation and Investment 

Decision Methods. Investment Evaluations Corp., 
Golden, Colo., 1974, 350 pp. 

24. U.S. Bureau of Mines. The United States Position and 

Outlook in Potash, BuMines IC 8487, 1970, 47 pp. 

25. Vine, J. D. Surface Geology of the Nash Draw Quad- 

rangle, Eddy County, N.Mex. U. S. Geol. Survey Bull. 
1141-B, 1963,46 pp. 



80 



APPENDIX A.— Assumptions and Estimations Used for the Economic Analysis 
of the Hypothetical Mining Unit B-1 (MU B-1) 



The operation is treated (for economic anal- 
ysis) as a separate company, not a subsidiary 
where losses could possibly be used for tax cred- 
its. A system diagram, showing approximate 
materials flow and gross revenue, and a com- 
puter printout of the investments and revenues 
analysis are included. 

It is assumed that two operating, competing 
companies in the Carlsbad area would form a 
third company to mine and process (through a 
mine complex and langbeinite wash plant) the 
ore determined to be in the area, provided a 
reasonable return on this investment could be 
expected from the project. A portion of this ore 
is located under the WIPP site. Sufficient ore 
exists outside the WIPP site to support a com- 
mercial potash operation, so that only the ore 
within the WIPP boundary is charged to the 
WIPP project. 

A langbeinite wash plant could be constructed 
at the mine site, and langbeinite product sold 
at the plant site; langbeinite fines produced 
would be shipped to two existing sulfate plants. 

Sylvite required at one of the sulfate plants 
(Sulfate Plant No. 2, fig. A-1 ) would be provided 
from its ongoing sylvite processing operation. 
MU B-1 is charged for mining and processing 
costs for sylvite fines produced in the normal 
processing operation. Sulfate processing costs 
are charged to and the sulfate revenues are cred- 
ited to MU B-1. 

Sylvite for the other sulfate plant (Sulfate 
Plant No. 1, fig. A-1) is provided from a mixed 
ore. The amounts of sylvite and langbeinite 
from the ore, plus the langbeinite fines from 
MU B-1, are estimated to provide capacity op- 
eration of this sulfate plant. This amount of syl- 
vite required results in a small excess amount 
of langbeinite from the mixed ore. Mining and 
processing costs for the mixed ore are charged 
to and the sulfate and excess langbeinite reve- 
nues are credited to MU B-1. 

Capital requirements include construction of 
(1) a langbeinite wash plant at the mine site; (2) 
the necessary storage and loading facilities to 
handle the langbeinite product and to ship the 
langbeinite fines produced to two existing sul- 



fate plants; (3) a mine including two shafts, un- 
derground facilities, and surface structures; and 
(4) reinvestment of some shorter life equipment 
in the sylvite ore, mixed ore, and sulfate plants. 
The capacity of Mining Unit B-1 is estimated 
to produce langbeinite fines in an amount ap- 
proximately equal to the present langbeinite 
fines feed to the existing sulfate plants. 

The ore deposits within this study area are 
estimated to contain a total of 79.78 million short 
tons (72.37 metric tons). In addition, an un- 
known amount of recoverable ore is estimated 
to occur outside the study area, contiguous to 
the deposits in the area. Of those deposits within 
the study area, 65.25 million short tons (59.19 
metric tons) would be required for MU B-1, of 
which 48.46 million short tons (43.96 metric 
tons) are within the WIPP site. The 79.78 million 
short tons (72.37 metric tons) of recoverable ore 
are in a minimum thickness of 4.5 feet, contain 
langbeinite in an average grade of 9. 1 1 percent 
KgO equivalent, and exist within reasonable 
mining distance of the approximate location of 
a production shaft. The 65.25 million short tons 
(59.19 metric tons) of ore deposits were used for 
the MU B-1 economic analysis. 

It is assumed the construction of the WIPP, 
with the consequent unavailability of the 48.46 
million short tons (43.96 million metric tons) 
that exist within the WIPP site, would not make 
the recoverable ore (outside the WIPP site and 
within the study area) uneconomic. Discounting 
the cash flows yields a rate of return on the 
estimated investment in excess of 15 percent. 
For this reason, a lease bonus bid amount (ac- 
quisition cost) was added to the other total initial 
investments. The amount of the lease bonus bid 
was determined so that, when added to all other 
investments, the project will provide a return of 
15 percent on the total investment. The amount 
of the lease bonus bid is also termed a present 
value. In other words, the producers could make 
an acceptable profit, arbitrarily assumed to be 
15 percent on their total investment, when that 
investment includes a sizable bonus bid amount 
for leases in addition to all other investments 
required for MU B-1. 



81 






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82 



The hypothetical MU B-1 plant is designed 
to handle a capacity of ore equivalent to 5,959 
short tons per day (5,406 metric tons per day) 
365 days per year. The average grade of the ore 
is 9. 1 1 percent K,,0 as langBeinite. The wash 
plant has an overall recovery of 87 percent. The 
plant produces about 21 percent fines; the bal- 
ance (89 percent) provides about 1,696 short 
tons per day ( 1 ,539 metric tons per day) of lang- 
beinite marketable product which, at $35/short 
ton, is $59,360 gross revenue per day. The hy- 
pothetical operation then produces 450.85 short 
tons (409.00 metric tons) of langbeinite fines for 
processing in the two sulfate plants. 

It is further assumed that about 179.0 short 
tons (162.4 metric tons), or 40 percent of the 
450.8 short tons (409.0 metric tons) of lang- 
beinite fines, is shipped to Sulfate Plant No. 1, 
where the sylvite-langbeinite ratio required for 
sulfate production is about 0.8927, indicating 
that 159.8 short tons (145.0 metric tons) of syl- 
vite is required to combine with the langbeinite 
for sulfate production. This sylvite is provided, 
at mining and processing costs, from the current 
mining operation at the sulfate plant. With a 
sulfate plant efficiency of about 80.6 percent, 
this provides about 218 short tons per day (198 
metric tons per day) of sulfate. At $88.50/short 
ton, this provides $19,293 gross revenue per 
day. 

The balance of the langbeinite fines from the 
MU B-1 plant are shipped to Sulfate Plant No. 
2, where the sylvite fines are provided from a 
mixed ore with a grade of about 9.2 percent K.^O 
as sylvite and 2.6 percent KgO as langbeinite. 
Mining Unit B-1 is charged mining and proc- 



essing costs for sufficient mixed sylvite and lang- 
beinite ore (the sylvite is combined with the 
langbeinite fines from MU B-1) for capacity 
operation of that sulfate plant. This results in 
a small excess of langbeinite, the gross revenue 
of which is credited to MU B-1. 

The 3,997 short tons per day (3,626 metric 
tons per day) of mixed fines normally produced 
in processing 9,608 short tons per day (8,716 
metric tons per day) of ore at 9.2 percent KjO 
sylvite grade with a sylvite recovery of 71.9 per- 
cent, would provide 440.68 short tons per day 
(399.78 metric tons per day) of sylvite fines. The 
3,997 short tons (3,636 metric tons) of ore is 
considered to be sylvite fines only, for the pur- 
pose of determining the costs to be charged to 
MU B-1. At 2.6 percent grade, 283.43 short 
tons (257.12 metric tons) of langbeinite is pro- 
duced with a langbeinite processing recovery of 
about 60 percent; 230.29 short tons (209.91 
metric tons) of this langbeinite is combined with 
the 271.85 short tons (246.62 metric tons) lang- 
beinite fines from MU B-1. These combined 
tonnages result in the production of about 
519.56 short tons (471.33 metric tons) of sulfate 
per day from this plant, which has an overall 
recovery of 69.3 percent. At $88.50/short ton, 
this provides a gross revenue of $45,981 per 
day. 

In addition, the 53. 1 3 short tons (48.20 metric 
tons) of langbeinite from mixed ore not needed 
in sulfate production are considered a credit, 
the "cost" charged to sylvite fines production. 
At $35/short ton, this provides a gross revenue 
of $1,859 per day. 



83 





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94 U. S. GOVERNMENT PRINTING OFFICE : 1980 - 31'+-720 






OEWUmiENT OF THE INTERIOR 
UNITED STATES GCOUJGKAL SURVEY 



EXPLANATION 



U.S. Geological Survey 
,. OPEN FILE REPORT 

ff r; ':: U .. G::io:icI"s:rvrs::nda.a. o^ „o-e„cUt«e. 

Eddy Co., N. Mex. 
ill 3,435 



Richardson and Boss 
Federoi-Welch No. I 

S«c. 20, T. 20 S., R. 31 E. 

Eddy Co., N Mex 

Alt. 3,524' 



OPEN FILE~1972 
BROKAW AND OTHERS 
CARLSBAD POTASH, 
NEW MEXICO 



v7/m 



wm 




f Salado Formation. 



liiiiiiiMii& r'-'' ' 

LIBRARY OF CONGRESS 






002 959 998 1 




